Zinc Flotation Concentrate Roasting Furnace

This paper describes experiments carried on at the Case School of Applied Science, together with their results. Their success led to the design of the larger furnace herein described, but which has not been built.

A previous article by the authors contained a general description of the new roasting furnace herein described but it did not go into detail as to the metallurgical behavior or the results obtained. Believing that such information would be of great value, they have elaborated on the subject and have given many unpublished details.

The furnace described applies the principle of roasting finely divided zinc-sulfide ores, now produced in large quantity by the flotation process, in gaseous suspension; that is, the ore particles are carried, in suspension, in a current of air and gaseous products of the roast. The relatively great fineness of flotation concentrate presents difficulties and problems of roasting in furnaces of the ordinary type; the fine ore is forwarded through the furnace in the form of a shallow bed and its very fineness leads to dense impervious bedding which prevents oxygen from reaching the interior of the bed, thus unduly lengthening the time of roasting and preventing the elimination of the last of the sulfur. The fineness of the concentrate, normally, should lead to a rapid and complete roast, for the speed of roasting is a function of the surface exposed to oxygen, which surface is greatest, per unit of weight, in very fine material. The difficulty in bed roasting is to get the oxygen to the particle. If the fine ore, during the roasting, could be freely suspended in oxidizing gases, full advantage could be taken of the great surface conferred by its fineness. This fundamental idea, of course, is not new, for the Stedtefelt furnace, familiar to the older metallurgists, is an example of it; but the manner in which this is accomplished may be new.

The numerous efforts to roast in gaseous suspension show that the idea is attractive; in fact, an analysis would indicate that it is the most reasonable way to effect the oxidation of ore, provided certain difficulties can be overcome. Two objections against such a method that formerly had much weight were the cost of fine grinding the ore, also the fact that the finely ground product, even after roasting, was not the best condition of material for further metallurgical operations. The objection of costly grinding, however, has been removed by the production of great quantities of flotation concentrate, which in point of fineness present a material that is ideal for roasting in gaseous suspension.

The inception of the experiments described here is due to David B. Jones and March F. Chase. The idea of the general type of furnace and process was suggested to the authors and the experimental work was carried out in the metallurgical laboratories of the Case School of Applied Science at Cleveland in 1915-16. The original plan was to make a furnace for roasting zinc-blende flotation concentrate that would use the heat value of the sulfide to accomplish the roast and to produce a gas suitable for making sulfuric acid; i. e., of sufficient concentration in SO2 and practically free from the products of carbonaceous combustion. Autoroasting of sphalerite is theoretically possible, and also practically, as was demonstrated in the experiments set forth.

A diagrammatic drawing of the roasting furnace and accessory apparatus as erected is shown in Fig. 1. Here A is the furnace proper; B the two stoves, heated by natural gas, that preheat the air; and C, the Cottrell electric precipitator for the precipitation of flue dust and fume carried from the furnace by the gases, which in a commercial plant would pass to a sulfuric-acid plant. At H is the gas supply for the stoves; X is the dust chamber, and G is the cycloidal blower that furnishes air to carry the ore in suspension in the furnace. The dried, preheated ore (60° to 100° C.) is charged into the hopper 1 whence the endless screw 2, operated by a variable-speed electric motor 3, discharges it into a pipe 4, directly above the nozzle 5. A stream of high-pressure (20 to 60 lb.) moderately pre-heated air is discharged through this nozzle in such quantity as to carry readily the fine ore in suspension into pipe 7, on the injector principle. Pipe 7 is of larger diameter than pipe 4 and is lined with refractory material; it serves as the main injector pipe into the furnace. Air, heated to approximately 800° C., passes from one of the stoves B to the injector pipe 7, through the supply pipe 9; this is the main air supply for roasting the ore.

The amount of air supplied is governed by two principles: (1) The quantity must be correct to roast the ore and to furnish a gas of the correct composition for the manufacture of sulfuric acid. Every pound of

sphalerite requires 35.7 cu. ft. of air (standard conditions) to convert the zinc to oxide, the sulfur to dioxide, with enough more oxygen to convert this into the trioxide. Some excess must be carried; in the experiments from 41 to 55 cu. ft. and sometimes 75 to 100 cu. ft. were used, as measured by a meter. (2) There must be such a relation between the quantity of air per minute and the area of the riser tube 10, that the velocity of the ascending air current will be enough to carry the largest ore particle to the top of this tube and over the edge. Fig. 2 gives definite data on this point for sphalerite. Fortunately the requirements for both conditions can readily be fulfilled.

The mixture of ore and high-pressure air from pipe 4 enters the main injector pipe 7 with a rotary motion and is caught by the ascending hot

air from the pipe 9 and injected into the riser tube 10. The fundamental idea is to have the temperature of the ore-air mixture at the ignition point of zinc sulfide, which varies between 650° and 810° C., depending on the size of the particle as it leaves the injector tube to enter the riser tube, so that the full time of the ore particle in the furnace will be available for oxidation. After leaving the riser tube 10, the ore particle falls in the annular space 11, the area of which is large so that the natural velocity of fall shall not be augmented by the velocity of the descending gases. It is the belief of the authors that the gas envelope surrounding the ore particle is constantly changing; thereby causing fresh oxygen to be supplied to the particle.

The time necessary for the roast is furnished by the passage of the ore up the combustion tube 10 and its fall in the annular space 11. This time depends on the height of the furnace, the length of the path of travel being practically twice’ the height of the furnace, also on the velocity of the ascending air current in the combustion tube. Assuming a definite ratio of cubic feet of air per pound of ore, the velocity will be determined by the area of the combustion tube. Definite figures on this point are given later. In any given furnace, i. e. a fixed area of combustion tube, there is a certain variation allowable in the velocity of the ascending gas current, obtained by varying the amount of air, which will then vary the time of the ore in the combustion tube. Too high a velocity, obtained either by too much air only or too great an ore feed with its corresponding increased amount of air, will shorten the time element so much that the sulfur is not sufficiently eliminated. The greater part of the ore collects in the hopper 14; the gases and fine dust pass, through openings 12, into the flue 13, thence to the settling chamber X, where the coarser dust is settled out, thence to the Cottrell precipitator C for the precipitation of the finest dust and fume.

Thirty-Foot Experimental Furnace

The design of the 30-ft. experimental furnace was, as far as possible, based on the results obtained in the previous work but was limited by lack of space and the capacity of the stoves, blowers, gas supply, etc. The stoves had been erected for the small furnace already described and were known to be inadequate for the larger furnace, but no space was available for enlarging them. It was also desirable to keep the expense as low as possible, so that some things that would have aided in the work were omitted.

The furnace consisted essentially of a brick stack (Fig. 3) approximately 30 ft. in height from the base plate to the top of the cover arch. The base plate and support rested on 4-ft. reinforced concrete posts, thus making the total height about 35 ft. The space between the posts provided room for the calcine hoppers and air and feed inlets and a pit provided the additional space necessary for the cleaning of tools and for making repairs.

The stack was built of two concentric firebrick circles 24 and 36 in. internal diameter; the space between these circles was filled with mineral wool. Somewhat larger circle brick were used near the bottom of the stack to increase the stability. No steel shell was placed outside of the brick, though such a shell would have aided greatly in the construction

of the furnace and in carrying on the tests. The top of the stack was closed by fireclay slabs, a small circular hollow tile being set at the center; a gate at its top permitted the interior of the furnace to be watched while operating. Two hoppers at the bottom of the stack collected the greater part of the roasted ore. Above these hoppers, and spaced at equal intervals around the stack, were four openings to the flue that led to the settler and to a small Cottrell precipitator.

Concentrically with the walls of the stack was placed the riser, or combustion tube, which extended from below the hoppers to within about 5 ft. of the furnace top. This tube, in the earlier trials, was made of

9-in. hexagon stove tile but later 12-in. cylindrical tile was used. The annular space between the combustion tube and the stack walls served as a downtake for the products of the roast.

The injector nozzle entered the bottom of the combustion tube for about 5 ft., thus placing its delivery point about 2½ ft. above the base plate of the furnace. The nozzle was a 2-in. fireclay tube and connected at its lower end with a pipe or mixer, into which the air and ore were injected from the stoves and feeder respectively.


The two stoves were of the central combustion type, the combustion tube being made of 9-in. hexagonal stove tile. The checkers were made of 2 by 2 by 9-in. brick laid to provide the maximum heating surface. A 1/8-in. steel shell with top and base plates insured airtight conditions. The stoves were 4 ft. in diameter, and 10 ft. in height. The usual burner, cold-blast, hot-blast, and flue openings were provided; the stoves were heated with natural gas.

Hot-blast mains connecting the stoves were 6-in. iron pipe lined with 4-in. fireclay flue liners, and covered with heavy asbestos insulation. Connections were provided between this pipe and the cold-air supply so that the temperature of the air entering the furnace was always under control. A bypass was provided at one point so that the air could be sent through a rotary gas meter at intervals; the metering was done cold and corrections made to temperature and pressure.


The feeder used was a modified Dunn pulverized coal feeder. Certain parts, originally made of brass, were later replaced by high-speed, hardened, tool-steel parts to avoid excessive wear from abrasion by the ore. A screw conveyor carried the; pulverized ore from the bottom of a hopper to a pipe in which a suction was produced by a compressed-air jet issuing from a 1/8-in. or 3/16-in. nozzle past its lower end. The ore was carried, by an expanding compressed-air current, toward the mixer and injector pipe. The connection into the mixer pipe was made tangentially and also inclined upwards so that the ore and air mixture met the hot air from the stoves in a rising spiral and thus thoroughly mixed the hot air and ore before they entered the combustion tube.

Approximately 10 cu. ft. of free air compressed to 60 lb. per sq. in. were necessary per pound of ore to carry the ore from in front of the 1/8-in. or 3/16-in. nozzle into the mixer. At times, even this amount was insufficient to prevent stoppage of the pipe, so the amount of air was increased at intervals to remove any accumulations by the arrangement shown in Fig. 5.

Seven thermocouples were placed in the furnace: four being placed in the combustion tube as shown, one in the mixer just below the base of the injecting nozzle, one in the downtake at the exit to the flue, and one in the downtake 20 ft. from the base plate. Lead wires were run to a central galvanometer station where readings were taken throughout each test.

General samples from the products accumulated during the run showed these sulfur contents:

This run began 1 hr. before that indicated (9:55) but conditions did not permit sampling, and adjustments caused unreliable results previous to those shown.

Total ore fed during the entire run, 2080 lb.; of this 160 lb. were fed before beginning the record.


Openings for oil burners were provided at intervals of 8 ft. throughout the height of the furnace. These openings were placed tangent to the inside circumference of the downtake. The burners were for the purpose of bringing the furnace to the operating temperature and were not used after such a temperature was reached and feeding of ore had begun.

The products of combustion of the oil heating were taken off at the top until the burners were going well, when the top was closed and the draft directed through the flues near the bottom of the stack. After a temperature of 800° to 1000° C. in the furnace had been reached, the burners were shut off, the burner openings closed, and the ore and air feed started. Adjustments of ore and air feed were then made as indicated necessary by the SO2 content of the gas, temperature conditions, and quality of roasted product. The furnace responded readily to such adjustments, but the time permitted for the test was not sufficient to permit adjustments to the best results possible.


In order to increase the capacity, the 9-in. tube was replaced by a 12-in. This tube was of specially made tile with 1½-in. walls, which restricted the area of the downtake more than was desirable and probably resulted in slightly poorer results than could have been obtained if the diameter of the furnace had been increased in proportion to the increase in the diameter of the combustion tube. Such a change, of course, was impossible without completely rebuilding the furnace.

Table 1 gives the results of a test made with the 12-in. tube. The temperature at the top of the furnace increased beyond that which was desirable, so in the last run cold air was admitted at 20 ft. from the base plate, thus providing for a complete control of the temperature and preventing the hot top. When cold air was thus admitted, it was necessary to reduce the amount of air fed with the ore, which meant a lowering of the velocity in the combustion tube and resulted in an increase in the bottom temperature. This increased bottom temperature, in turn, permitted the reduction in the temperature of the air coming from the stoves.

It was thought that accretions would form in the combustion tube during the roasting but, with one exception, no trouble was encountered from this source, even though the temperatures at times were allowed to go beyond those that would be permitted in practice. Some little difficulty developed at points near the oil burners, as in preheating the temperature in these regions was necessarily higher than the intermediate zones. The temperature of the various zones, however, were quickly equalized and no trouble was encountered during the first few tests in the furnace, the combustion tube being clean at the end of the test; in later tests, however, after some dust had collected in crevices of the tube, the overheating at the burners caused some slagging between the dust and fireclay, thus developing starting points for accretions. It is not thought that this would cause any trouble in practice for, after the furnace is in operation, it will continue so for a considerable time. In this experimental work, the operations were necessarily short and at infrequent intervals, which meant repeated heating of the furnace with consequent trouble from slagging.

The gases and fine dust were taken from the furnace into the flues through four openings equally spaced around the base of the stack. The flue led first into a settling box, in which the rate of flow of the gases was reduced and screens prevented the channeling of the currents. From this settler, the bases passed through a small Cottrell precipitator for final cleaning. The cleaning was very effective, though this end of the operation was given but little attention. The Cottrell equipment contained four 10-in. tubes 16-ft. long which were thoroughly grounded. In the center of these 10-in. pipes, and connected with the 50,000-volt line, were suspended, on insulators, ½-in. pipes fitted with four knife edges at right angles to one another. The 400-volt a.c. current was stepped up to 50,000 volts and rectified by a Kenetron.

The flotation concentrate used in the furnace was a complex sulfide containing 31.4 per cent, sulfur, 44.3 per cent, zinc, 11.6 per cent. iron. The results of the roasting of this material and conditions existing during the roast are given in Table 1. These show a minimum sulfur content in the roasted product of a 2.2 per cent, and 3.6 per cent, for an afternoon period after the furnace had been adjusted to best running conditions. Other tests gave continuous results of slightly over 2 per cent, sulfur.

Tables 2, 3, and 4 show the forms in which the sulfur existed in the roasted product. The particular point of interest in these results is that the sulfur existed largely as sulfate and the quantity as sulfide was small. Table 2 shows the per cent, of SO2 in the gas corresponding to the roasted ore samples. It is evident that the concentration of gas had no effect in preventing the completeness of the roast; i. e., the sulfur is not higher in roasted products produced in higher SO2 concentration.

Note.—Decreased sulfur content of calcines in the afternoon results from the higher furnace temperatures, as shown by Table 1.

Note.—During this run the SO2 content of the gas in the furnace was fairly constant at 9 to 9.5 per cent, during the greater part of the run.

Hopper sample is calcine collected from the hoppers of the furnace; flue sample is calcine cleaned from the horizontal flue leading to the settler; settler sample is calcine taken from the settler; precipitate or sample is calcine taken from the Cottrell precipitator. The coarsest material was from the hopper and the finest from the precipitator.

Tables 3 and 4 show that the sulfur exists in the roasted product as sulfide and sulfate, and that the sulfate sulfur is predominant in the fine ore. This is shown not only by the general samples, of which the hopper samples are the coarsest and the precipitator the finest, but also by the screen analysis of the roasted products. High sulfide sulfur is present only in coarse material, which in total amount is insignificant in flotation concentrate. The total sulfur is highest, in general, in the finest product, but consists largely of sulfate sulfur. Work that cannot be detailed here reveals that this sulfate sulfur is not the result of an incomplete roast m the furnace, but results from a resulfatization of the roasted ore. This resulfatization is due to relation between partial pressures of SO2 and SO3 and the temperature existing in the lower part of the furnace, in the settler, and in the precipitator. Generally, the partial pressure of the sulfur gases is such that no resulfatization of the calcine can take place if the furnace temperature is between 900° and 1000° C. But when this temperature falls to near 800° and below at the flues, the rapid resulfatization of the ore commences. For example, it will be noted that in Tables 3 and 4, the minus 200-mesh material of the hopper samples, i. e., material separated from the furnace at high temperature and immediately cooled, is much lower in total sulfur, as well as sulfate sulfur, than material of the same size from the flue, settler, or precipitator samples, which were subject to the lowering critical temperatures in the presence of sulfating gas. When the aim of the roasting is a practically complete desulfurization, the furnace must be run hot and the calcines and gas rapidly cooled, or separated from each other as soon as they leave the furnace. If the roasting is carried out as a preliminary for sulfuric-acid leaching, as in the electrolytic-zinc process, the sulfating action of the furnace may be used to any desirable degree up to its maximum.

The Proposed East St. Louis Furnace

An increase in height was provided for in the design of the 12-ton furnace it was planned to build at East St. Louis, but which was not completed because of the conditions brought about by the war. Fig. 6 shows that the combustion tube is set to one side instead of concentrically with the downtake. With the concentric combustion tube, it was necessary to place the furnace on posts to provide space for feed and air connections and for repairs. It was also difficult to make pyrometer and air connections from the outside to the combustion tube, at intervals throughout the height, and to provide support for the combustion tube. To overcome these objections, the side combustion tube was designed.

Air and Feed Arrangement

Provision was made for bringing all the air for the roasting from the stoves through the mixer, as in the smaller furnace, or for splitting the current and having a part of the air take this path and the remainder enter the combustion tube near the top of the injector nozzle. Pipe connections to hot- and cold-air sources were placed at intervals along the combustion tube, thus providing for an absolute control of the

temperature in this tube; Fig. 7 illustrates this arrangement of the ore and air feed connections.


As in the smaller furnace, a screw conveyor takes the ore from the bottom of the feed hopper to a pipe immediately above a compressed-air

injector. The injector was changed considerably from that of the smaller feeder, in order to reduce the amount of compressed air necessary for injection purposes. As shown, the compressed-air nozzle blows directly
through a throat, or Eureka siphon, carrying the ore with it and delivering it into a current of low-pressure air coming from hot or cold sources, or both, by which it is carried forward into the mixer.

Experiments indicate that not to exceed 20 cu. ft. of free air compressed to 30 lb. will be necessary for injecting 1 lb. of ore as against 10 cu. ft. of free air compressed to 60 lb. in the feeder used on the small furnace.

The delivery pipe from the feeder was to be connected to the mixer pipe by the tangential and rising connection that was used on the smaller furnace. The same principle was also to be used in connecting the excess-air pipe into the combustion tube.


The bottom, or base, plate of the combustion tube was to be raised about 10 ft. from the floor line, to provide for feeder and air connections. The space opposite these connections in the base of the downtake was to be used for the settling chambers for the roasted ore. Accordingly provision was made for the removal of the coarse product from above a bulkhead in the downtake at a point opposite the injector nozzle. The fines and gas were to be taken off, somewhat above this, through three radially placed flues, which flues in turn entered tangentially a settling chamber placed beneath the bulkhead just mentioned. This method of entry was to provide a circular or centrifugal motion for settling the dust. The gases and remaining dust were to be taken off through a flue passing upward and outward through the bulkhead, thence to the Cottrell precipitator.

Heating the Air

There is insufficient data as to how much of the air must be heated and to what temperature heating should be carried for the most efficient operation of the furnace. The effect of heated air on the behavior of the furnace and the ignition has been discussed and it is obvious that one of four methods must be used.

  1. Heating all the air and passing all of it through the mixer tube and into the furnace with the ore; this will require large stove construction and will not allow air changes in various parts of the furnace, as will be desirable for proper control.
  2. Heating all the air and passing only a part of it into the furnace through the mixer, the remainder to be introduced into the combustion tube at various points; this method is not desirable as the heat is needed in the bottom of the furnace for. heating the ore and bringing about early ignition.
  3. Heating none of the air; indications are that this will not be possible, as already explained.
  4. Heating a part of the air (say one-fifth), which will be passed into the furnace with the ore through the mixer, the remainder of the air to be admitted cold at points throughout the height of the furnace; this method will make all the stove heat available for ignition and for holding down the combustion zone, and will permit the use of cold air for combustion and temperature control of the top of the furnace. Of course, the stoves should be of ample capacity and of such construction as to permit changing the temperature and amount of air going into any part of the furnace as may be found necessary.

Under the first condition, assuming that 300° C. is necessary for mixer temperature, the air in a 50-ton furnace will have to be heated to 350°
in order to heat the ore, if it is being fed cold. If the ore is taken direct
from the dryer, this temperature may be reduced somewhat.

The authors calculate that three stoves of the following dimensions will readily take care of a 50-ton furnace: 1060 sq. ft. heating surface, 246 flues, 4 by 4 in. in cross-section, 14 ft. long, and 3¼-in. walls; total height allowing for gas and air ports, foundation and dome about 25 ft.; diameter about 12 ft.;.cross-section of combustion chamber 12 sq. ft.

Under the fourth condition, which seems the most rational, such stoves would be more than ample, but not necessarily inefficient, and would provide for emergencies where the heating of all the air was necessary. The ratio of fuel necessary for heating all the air to 250° to that necessary for heating one-fifth to 600° is about as 2.5 is to 1, which is quite a saving in favor of the fractional heating.

Drying the Ore

It is obvious that the feed for this type of furnace must be dry, or it will not pass the feeder and injector. It is likely that most of the concentrate delivered at the roasting plant would have to be dried, so it was planned to do this drying with the waste heat in the gases leaving the furnace, by circulating them beneath the hearth of an ordinary kiln dryer.

That there is sufficient waste heat available for this purpose is readily shown. In the experimental furnace, the gas left the furnace at 800° C. under conditions of best roasting. The volume of this gas (measured under standard conditions) was approximately 50 cu. ft. per lb. of ore roasted and was made up of about 8 per cent, sulfur dioxide; 10 per cent, oxygen, and 82 per cent, nitrogen.

Assuming that these conditions are obtained in a 50-ton furnace:

50 tons per day = 69.5 lb. or 31.5 kg. per min.
69.5 lb. will produce 3475 cu. ft. per min. of gas.
8.0 per cent, of 3475 = 278 cu. ft. SO2 = 22.7 kg. per min.
10.0 per cent, of 3475 = 347.5 cu. ft. O2 = 14.1 kg. per min.
82.0 per cent, of 3475 = 2849.5 cu. ft. N2 = 101.7 kg. per min.

Further, assuming that the gas is lowered in temperature by 400° in passing the drier, the heat available there is:

  1. Sulfur dioxide:
    Heat capacity at 800° C. is 1 X 800 (0.125 + 0.0001 X 800) = 164; 164 X 22.7 = 3722.8 kg.-cal.
    Heat capacity at 400° C.
    66 X 22.7 = 1498.2 kg.-cal.
    Heat available from SO2 = 2224.6 kg.-cal.
  2. Oxygen:
    Heat capacity at 800° C. per kg. = 181 cal.
    181 X 14.1 = 2552.1 kg.-cal.
    Heat capacity at 400° C. per kg. = 88 cal.
    88 X 14.1 = 1240.8 kg.-cal.
    Heat available from oxygen = 1311.3 kg.-cal.
  3. Nitrogen:
    Heat capacity at 800° C. per kg. = 207 cal.
    207 X 101.7 = 21051.9 kg.-cal.
    Heat capacity at 400° C. per kg. = 100 cal.
    100 X 101.7 = 10170.0 kg.-cal.
    Heat available from N2 = 10881.9 kg.-cal.
    Total heat available per minute;
    2224.6 + 1311.3 + 10881.9 = 14417.8 kg.-cal.

Assuming 20 per cent, moisture in the ore, it will require 39.3 kg. of moist ore to provide the 31.5 kg. of dry ore. This means that 7.8 kg. of water must be evaporated each minute, and 39.5 kg. of ore must be raised to 200° C.

It is, therefore, evident that the necessary heat is available for drying and heating the ore. If this heated ore is fed without cooling to the furnace, the temperature of the air from the stoves may be reduced somewhat from that necessary when cold ore is fed. As this amount of waste heat was actually delivered by the small experimental furnace, it is clear that enough heat would be available for completing the roast and maintaining furnace temperatures. As this was possible with a small furnace with a high radiation factor, it would be more readily accomplished in a larger well-insulated furnace.

The dried ore from the dryer will contain some lumps, formed in the drying, and some coarse particles that will not work well in the feeder; it was planned, therefore, to interpose a vibrating screen between the dryer and the feed hopper to remove these, insuring a satisfactory feed at all times.


The capacity of these furnaces will vary as the square of the diameter of the combustion tube. This is substantiated in the results from the 9-in. and 12-in. combustion tubes used in the experimental furnace, which had capacities of 2.0 and 3.6 tons per 24 hr., respectively. Assuming similar conditions as to ratio of air and ore, and the velocity in the combustion tube, the diameter necessary in the tube of a 50-ton furnace will be 45 in. It may be that for a furnace of this capacity it would be better to use two tubes having a combined area equal to 45 in. diameter, i. e., 32 in. diameter each. Such an arrangement would provide for repairs in feeders, nozzles, and injectors on a part of the furnace without complete shutting down the furnace, with attendant cooling.


Effect of Oxygen-enriched Air in Roasting Zinc Ores

Experiments have shown that the use of enriched air would be of particular benefit in the roasting of zinc ores for the manufacture of sulfuric acid. Enriched air increases capacity of furnace, decreases fuel consumption, and increases SO2 content of roaster gas.

The work here described had for its purpose the procuring of data from which some quantitative estimate might be made of the results obtainable by using oxygen-enriched air in roasting zinc ores on a commercial scale. The principal metallurgical advantages of using enriched air in roasting zinc ores would be:

  1. The rate of roasting would be increased, with consequent gain in the capacity of the roasting furnace.
  2. As less air would be required for roasting, the volume of hot gases leaving the furnace and the heat carried out of the furnace as sensible heat in these gases would be less per ton of ore roasted; partly for this reason and partly because of the increased quantity of heat generated in the furnace by the larger amount of ore that could be roasted, the consumption of fuel by the furnace would be lessened; and by the use of air sufficiently enriched with oxygen the necessity of using fuel might be entirely obviated.
  3. Roaster gas having a higher SO2 content could be produced; this would make possible greater capacity and more economical operation of the sulfuric-acid plant.

Certain phases of the application of enriched air to roasting can be worked out only by experimenting with a furnace of commercial, or at least semicommercial, size. Thus the precautions necessary to secure proper distribution of heat in the furnace; the volume of enriched air, and the proportion of oxygen in this air, necessary to give the desired increase in roasting capacity of the furnace and in SO2 content of the roaster gases; and the most desirable frequency of raking, thickness of ore bed, and rate of advance of the ore through the furnace can be definitely determined only after such large-scale experiments.

On the other hand, by drawing up suitable heat balances, the amount of additional heat, per ton of ore roasted, made available in the furnace by the use of enriched air can be calculated, and from that can be calculated, if it is assumed that proper distribution of the total heat in the furnace can be arranged, the additional amount of ore that must be roasted per unit of time in order to make the use of fuel unnecessary. Furthermore, data concerning the effect of enriched air on the ignition temperature and rate of oxidation of zinc ores can be obtained by means of laboratory experiments in which all other conditions (such as temperature, volume of air supplied, thickness of the ore bed, size of; the ore particles, and frequency of stirring) can be maintained constant. From these data, estimates may be made of the increased capacity that may be expected in a full-size furnace, and of the increase in SO2 content of the roaster gas that may be expected, as a result of the use of enriched air.

Heat Balances of a Hegeler Roaster Using Ordinary Air and Using Enriched Air

Heat Balance of a Hegeler Roaster Using Ordinary Air

When the suggested use of enriched air for roasting zinc ores was first called to the attention of the writers, they drew up a heat balance of a Hegeler roaster using ordinary air; this particular type of furnace was selected because it is the furnace generally used in this country for roasting zinc ores when the gases are to be utilized for making sulfuric acid. The heat balances of different Hegeler roasting furnaces will vary in detail, depending on the design of the furnace, composition of ore roasted, kind of gas producers used, quality of coal used, and manner of preheating the air for roasting, but the variations are in the minor items; the net result, as shown by the consumption of coal per ton of ore, is nearly the same in most furnaces of this type.

The heat balance here given is for a hypothetical case, in that the operating, data were not taken from the actual operation of any one particular furnace. The conditions assumed were, however, representative of actual practice, so that the heat balance is typical. A few simplifying assumptions were made, such as assuming an ore consisting entirely of ZnS, FeS, and SiO2, minor constituents being neglected. The effect of such simplifying assumptions on the accuracy of the calculations is negligible.

This heat balance of a Hegeler roaster, using ordinary air and roasting 45 tons of 60 per cent, zinc ore per day, is summarized in Table 1.

Heat Balance of a Hegeler Roaster Using Enriched Air

In the heat balance of a Hegeler roaster using enriched air, if it be assumed that this use of enriched air is to eliminate the use of fuel, several of the items enumerated in Table 1 will be absent. These are, from the debit side, the sensible heat in the preheated air (as there will be no waste combustion gases for preheating), the sensible heat in the producer gas, and the heat of combustion of the producer gas; and from the credit side, the sensible heat in the combustion gases. There remains then, as a source of heat in the furnace, only the oxidation of the ore, which must balance the heat lost as sensible heat in the roasted ore and in the roaster gases leaving the furnace, and that lost by radiation and conduction.

Assuming the composition of the ore before and after roasting, the temperature of the green ore and of the air supply entering the furnace, and the temperature of the roasted ore and roaster gases leaving the furnace, to be the same as they were assumed for the purpose of calculating the heat balance of the roaster using ordinary air, and assuming that the loss of heat from the furnace by radiation and conduction would remain constant, the tonnage of ore that would have to be roasted per 24 hr. to maintain the furnace at roasting temperature without the use of other fuel was calculated for the following cases:

Case 1.—The enriched-air supply to contain 25 per cent, oxygen; the exit gases to contain two volumes of SO2 to one volume of O2 (in this case 13.3 per cent. SO2 and 6.65 per.cent. O2); 60 per cent, of the roaster gases to be recirculated and returned to the furnace at 200° C. to help in controling the rate of combustion and the distribution of the heat in the furnace and in procuring the desired high content of SO2 in the gases. The flow sheet under these conditions is shown in Fig. 1.

Without going into details, the calculations may be summarized as follows: From the combustion of 1000 lb. of the ore are obtained 1,049,650 lb.-cal. The heat leaving the furnace as sensible heat in the roasted ore (853 lb.; see flow sheet) at 800° C. is 106,350 lb. cal. and in the roaster gases (63,140 cu. ft.) at 600° C., 775,888 lb.-cal.; of the latter, 142,644 lb.-cal. are returned to the furnace in the recirculated roaster gases (37,884 cu. ft.) at 200° C. Thus 1,049,650 – 106,350 – 775,888 + 142,644 = 310,056 lb.-cal. are available per 1000 lb. of ore roasted, to balance radiation and conduction losses.

From the previously calculated heat balance of a Hegeler roaster using ordinary air, it was found that the loss of heat from the furnace by radiation and conduction was 111,696,200 lb.-cal. per 24 hr. From this, it follows that 111,696,200/2 x 310,056 = 180 tons of green ore must be roasted in the furnace per 24 hr. to maintain it at the usual roasting temperature, using enriched air under the conditions assumed in this case.

Case 2.—Conditions the same as in Case 1, except that none of the roaster gases are recirculated; or, what amounts to the same thing thermally, that the recirculated gases are returned to the furnace at the same temperature as that at which they leave. Under these conditions the flow sheet is as shown in Fig. 2.

The heat obtained from the combustion of 1000 lb. of the ore is, as before, 1,049,650 lb.-cal., and the heat leaving the furnace as sensible heat in the roasted ore is 106,350 lb.-cal., the heat leaving the furnace as sensible heat in the roaster gases is, however, only 40 per cent, of what it was in the former case, or 310,355 lb.-cal. The heat available to balance radiation and conduction losses is then: 1,049,650 — 106,350 — 310,355 = 632,945 lb.-cal., and 111,696,200/2 x 632,945 = 88 tons green ore must be roasted per 24 hr. to maintain the furnace at the usual roasting temperature, using enriched air under the conditions assumed in this case.

Case 3.—The enriched air supply to contain 50 per cent, oxygen; the exit gases to contain two volumes of SO2 to one volume of O2, which in this case means that the SO2 content will be 28.6 per cent, and the O2 content 14.3 per cent.; none of the roaster gases to be recirculated. The flow sheet under these conditions is shown in Fig. 3.

In this case, the heat leaving the furnace as sensible heat in the roaster gases at 600° C. is, per 1000 lb. of ore roasted, only 155,178 lb.-cal. and the heat available to balance conduction and radiation losses is
1,049,650 – 106,350 – 155,178 =788,122lb.-cal. Therefore, 2×788 122 = 71 tons of green ore must be roasted per 24 hr. to maintain the furnace at the usual roasting temperature.

The SO2 content of the roaster gases in the examples just discussed was purposely assumed to be very high. If it should be impracticable to obtain such a high SO2 content in the roaster gases, a larger amount of ore would have to be roasted to produce the same amount of available heat in the furnace. Even with the SO2 content assumed as high as it has been, the increase in the amount of ore that must be roasted per 24 hr. in order to dispense with the use of file] is considerable.  It is open to question whether such high SO2 content in the roaster gases, with simultaneous large capacity of the furnace, could be accomplished except by the use of enriched air containing a very high percentage of oxygen.

On the other hand, it might seem possible that the ignition temperature of zinc blende in enriched air would be so much lower than in ordinary air that the roasting furnace would not have to be run at so high a temperature when enriched air is supplied as when only ordinary air is available; this would permit a saving in the heat lost by radiation and conduction, and as sensible heat in roaster gases leaving the furnace. To obtain experimental evidence bearing upon these questions, the series of experiments here described was undertaken.

Ignition Temperatures of Sphalerite in Air Enriched with Various Proportions of Oxygen

In giving data on ignition temperatures, it is necessary to define exactly what is meant by the term ignition temperature. The usually accepted meaning is the temperature at which the oxidation of a substance becomes so rapid that the heat liberated counterbalances the heat radiated or conducted away, thus maintaining a visible spontaneous combustion. The temperature at which this can take place varies with the rate at which heat is radiated or conducted away from the substance; this in turn is affected by the heat conductivity of the walls of the vessel in which the substance is contained and by the volume and temperature of air circulated over it. ”

Oxidation of sphalerite exposed to the air, no doubt, takes place at an extremely slow rate, even at ordinary atmospheric temperatures. It is conceivable, if a pile of finely divided zinc blende could be so insulated that radiation and conduction from the pile would be nil and the air supply so regulated that the heat carried off by it as sensible heat would be as small as possible, that the slow oxidation of the blende would, of itself, cause the pile to become sufficiently hot for active combustion to take place. This would be analogous to the spontaneous ignition of large coal piles. In such a case, it would be difficult to say just what should be called the ignition temperature.

In outlining the series of determinations of the ignition temperatures of sphalerite in air enriched with various proportions of oxygen, it was at first planned to heat slowly a sample of the mineral in an electrically heated tube, passing the enriched air over it at a fixed rate; to read the temperature of the sample at intervals by means of a thermocouple, the junction of which was placed in the sample; and then to plot a curve of the rate of temperature rise. It was thought that at the temperature of ignition there might be a sufficient increase in this rate to cause a noticeable deflection in the curve; it was found, however, that the oxidation of the sphalerite began so gradually that no such deflection could be detected.

It was then decided to determine the temperature at which sufficient sulfur dioxide was formed to cause to turn blue a solution of potassium iodate and starch placed at the exit of the tube containing the sample. The ignition temperature, even as determined by this method, varied according to the rate at which the sample was heated, the rate at which the air was passed over the sample, etc., but by heating very slowly and keeping the rate of heating and other variable factors the same in all the experiments, comparative results were obtained that show clearly the effect on ignition temperature of increasing the oxygen content of the air supply.

Apparatus and Procedure

The oxygen-enriched air for use in the experiments was made by mixing commercial oxygen with ordinary air in a large gas-storage bottle. The pressure in this storage bottle was regulated by raising or lowering a pressure bottle of the same size filled with water, which was placed on a small elevator and connected to the storage bottle by a flexible siphon. Before the gas was passed over the sphalerite; it was passed through two washing bottles containing, respectively, sodium-carbonate solution and distilled water, and through two drying tubes containing anhydrous calcium chloride.

The sample of sphalerite was placed in a pyrex glass combustion tube 20 mm. in diameter, which could be heated by means of a nichrome-wound electric-resistance furnace. There were two sections of this furnace, one of which was used to heat the sphalerite and the other to preheat the air so that it would have about the same temperature as the sphalerite before coming into contact with it. The temperature of the sphalerite was read by means of a thermocouple, the junction of which, protected by a thin quartz tube, was placed in contact with the surface of the sample. The temperature of the preheated air was read by means of a second thermocouple. One end of the combustion tube was connected to the supply of enriched air; the other (exit) end to a small washing bottle containing a few cubic centimeters of a solution of potassium iodate and starch, to serve as an indicator for sulfur dioxide. Beyond this bottle of indicator solution, there was attached a flow meter for measuring the flow of air through the system.

The sphalerite used was a hand-picked specimen of the massive mineral. It had the following analysis: zinc, 65.17 per cent.; sulfur, 32.36 per cent.; iron, 0.48 per cent.; insoluble, 0.79 per cent. As the size of the particles of the sphalerite has a marked effect on the ignition temperature, the crushed sample was separated by screening into four sizes: through 20, on 28 mesh; through 28, on 35 mesh; through 35, on 100 mesh; and through 100 mesh; and a separate series of experiments run on each size.

A 15-gm. sample of sphalerite was used for each experiment. It was placed in the combustion tube, the thermocouple placed in position, the gas train made tight, and enriched air of the desired oxygen content passed until the apparatus was filled with it. The gas flow was then adjusted to a rate of 13.5 liters per hour, which had been selected as a standard for the experiments. The current was turned on in the furnaces for heating the sphalerite and for preheating the air; these were heated rapidly up to 30° to 40° C. below the expected temperature of ignition and then at the rate of 1° C. per min. until the temperature of ignition was reached, as indicated by the potassium iodate-starch solution turning blue. The temperatures of the sphalerite and the preheated air were at all times held approximately the same.

The ignition temperatures as determined by the above method are tabulated in Table 2; Fig. 4 shows curves plotted from the values given in this table.


The results of these experiments show that the ignition temperature of sphalerite is appreciably lowered by increasing the oxygen content of the air supply. This lowering is, however, very, small, the ignition temperature in pure oxygen averaging less than 25° C. below that in ordinary air containing only 21 per cent, oxygen; therefore, the effect of enriched air on ignition temperature would be of very slight practical importance in the roasting of zinc ores.

Rates of Oxidation of Sphalerite in Air Enriched with Various Proportions of Oxygen

Apparatus and Procedure

The apparatus used for this series of experiments was similar to that used for the determination of ignition temperatures, which has been described, except that the bottle of potassium iodate-starch solution was omitted. The sphalerite used was also the same and, as before, separate experiments were run on the following sizes through 20, on 28 mesh; through 28, on 35 mesh; through 35, on 100 mesh; and through 100 mesh.

A 5-gm. sample of sphalerite was used for each experiment. It was spread over the bottom of an alundum boat in a layer about 1/8 in. thick, and was not stirred during the roasting. When starting an experiment, the apparatus was filled with air of the desired oxygen content. The furnace for heating the sample and that for preheating the air were then started and raised rapidly to a temperature of 750° C. The gas flow was adjusted to a rate of 5 liters per hour, and the temperature held constant at 750° C. for 1 hour. The furnace was then allowed to cool rapidly and the partly roasted sphalerite was analyzed for total sulfur and water-soluble sulfur.

The results obtained in the experiments are tabulated in Table 3, and plotted as a series of curves in Fig. 5.


Theoretically, other conditions being equal, the rate of oxidation of zinc blende should vary directly with the partial pressure of oxygen in the air to which it is exposed. The curves in Fig. 5 show that this is borne out fairly well by the experiments in which the oxygen content of the air supplied was less than 50 per cent. With higher concentrations of oxy¬gen, the elimination of sulfur did not increase in the same ratio as the oxygen content of the air. In these experiments with air of high oxygen content, however, the sulfur was reduced to such a low point in 1 hr. that the surface of the blende particles was, no doubt, much reduced and was covered with a coating of zinc oxide sufficient to retard the rate of oxidation decidedly. In the experiments with the -35- + 100-mesh, and the —100-mesh sphalerite, the sulfur elimination was less in pure oxygen than in 50 per cent, oxygen-air; also in 50 per cent, oxygen-air and in pure oxygen the sulfur elimination was less from the —100-mesh than from the -35- + 100-mesh size. This is explained by the fact that the finer sizes, when roasted in air of high oxygen content, tended to sinter and form a cake. No doubt the surface temperature of the sphalerite, because of the rapid oxidation in oxygen, was considerably higher than the temperature of the furnace.

It would probably be safe to state, as a result of these experiments, that under similar conditions, the rate of oxidation of sphalerite varies very nearly directly as the partial pressure of oxygen in the air to which it is exposed, at least for all concentrations of oxygen likely to be used in roasting on a large scale.

A second fact is the large amount of water-soluble sulfur in the calcine from roasting in air of high oxygen content. This indicates that the tendency to form zinc sulfate in the preliminary stages of roasting would be greater with enriched air than with ordinary air. It would be necessary to break these up in the final stage of roasting; this might require a higher temperature or a longer time at a high temperature at the end of the roast than present practice requires.

This series of experiments concerning the effect of enriched air on the rate of oxidation of sphalerite is incomplete. In the experiments just described, the sphalerite was roasted for a definite time in all the experiments, consequently in the experiments with the finer sizes and with enriched air of high oxygen content the sulfur in the sample was reduced to a much lower point than in the experiments with the coarser sizes and with air of lower oxygen content. The rate of oxidation naturally decreased as the sulfur content of the sample decreased, and this effect counterbalanced to a certain extent the effect that the increased oxygen content of the air had of increasing the rate of oxidation. It appears now that a better method of experiment would have been to roast all samples to the same content of sulfur and compare the time required to do this with enriched air of various oxygen contents. It was considered unnecessary, however, to carry this series of experiments any further, as more reliable information is given by the experiments next to be described, which were carried out with a roaster capable of taking a charge of several pounds of ore.

Experiments with a Mechanically Rabbled Laboratory Roasting Furnace

Apparatus and Procedure

The laboratory roasting furnace used in these experiments was an electrically heated, mechanically rabbled furnace, patterned after one used by C. A. Hansen for experiments in the roasting of zinc ores for leaching. It is shown in Fig. 6. A sheet-iron cylinder 30 in. in diameter and 18 in. high was set on timber skids, as a foundation, and a layer of heat-insulating brick laid in the bottom. A heavy sheet of iron was laid level on the layer of brick and on this, concentric with the outer sheet-iron cylinder, was set a thin cast-iron cylinder 16 in. in diameter and 9 in. high. Inside of this inner cylinder was laid a layer of firebrick, covered with about ½ in. of crushed firebrick. On this was set the heating unit, which was a fireclay disk with shallow grooves running transversely across the upper surface in which was wound the heavy chromel wire that served as a resistor. On top of the heating unit a thin fireclay disk was placed. Above this hearth bottom the cast-iron cylinder was lined with a fireclay cylinder ¾ in. thick. On top of the cylinder rested a sheet of asbestos and a heavy iron plate, with a hole in the center for the shaft to which the rabble arms were keyed. All joints in the furnace lining were sealed with alundum cement. An opening 4½ in. wide by 3½ in. high was left in one side of the furnace as a door; it was closed with a firebrick plug.

The rabble arm and rabbles were formed from a single piece of heavy strap iron. The rabbles were so arranged that the ore was thoroughly stirred and, at the same time, maintained at uniform depth over all the hearth. The rabble arm was driven by a small motor and worm gears at a speed of 0.95 r.p.m. The driving mechanism for the rabble arm was supported on a slab of hard asbestos board resting on top of the furnace.

The space between the inner cylinder, forming the roasting furnace proper, and the outer sheet-iron cylinder, was filled with infusorial earth for heat insulation.

The shaft carrying the rabble arm was hollow, and a carefully calibrated platinum-platinum rhodium thermocouple, with silica protecting tube, was inserted through it so that the end rested upon the floor of the roasting hearth. The power input to the furnace was controlled by a voltage regulator; in this way the temperature of the furnace could be regulated to within ±10° C. The temperature of the roasting ore was difficult to determine accurately; occasional readings, taken with a ther-

mocouple thrust into the layer of ore while the rabble arm was stopped, averaged about 20° C. higher than the temperatures read with the thermocouple in the central shaft.

Air for roasting, either atmospheric or enriched, as the case might be, was admitted to the furnace through a ½-in. pipe, curved to direct the incoming air away from the gas outlet and sample tube. The roaster gas left the furnace chiefly through the small cracks around the plug that was inserted in the door of the furnace. Samples of the gas for analysis were drawn off through a silica tube not far from the gas outlet.

Sulfur dioxide in the roaster gases was determined by absorption in potassium-hydroxide solution, and oxygen in the roaster gases and in the air supply by absorption in alkaline potassium-pyrogallate solution, in an Orsat apparatus.

The method of controlling the volume and composition of the air supply is shown in Fig. 7. The atmospheric air was supplied by a small

laboratory blower; a large glass carboy was placed in series with this to act as an accumulator to diminish the effect of minor fluctuations in pressure. Oxygen was supplied from a cylinder of the compressed gas. Flow meters were inserted in both the air and the oxygen supply lines to indicate directly the rates of flow of air and oxygen. By maintaining a constant reading on each of these flow meters, the volume and composition of the air supplied to the furnace could be maintained constant within about 1 per cent. In series with the flow meters were wet gas meters, serving as integrating meters on which could be read the total volume of gas passed in any given interval of time. The oxygen and air supply lines led into a large glass bottle, which served as a mixing chamber; the outlet from this bottle was fitted with a three-way stop cock, of which one outlet led to the roasting furnace and the other to the gas analysis apparatus. The complete equipment is shown in Fig. 8.

The ore used was Joplin concentrate, screened through a 10-mesh screen to give a product of fairly uniform size. Its composition was: zinc, 62.07 per cent.; lead, 1.29 per cent.; iron, 1.41 per cent.; sulfur, 31.58 per cent.; insoluble, 1.84 per cent.; CaCO3, 0.59 per cent. Its

screen analysis is given in Table 4. For each experiment, 7 lb. of this ore was used; this made a layer in the furnace about ¾ in. thick.

When starting an experiment, the furnace was heated to the temperature at which the experiment was to be run the air, of the desired oxygen content, was turned into the furnace; and the charge of ore was placed in the furnace and spread evenly over the hearth. The introduction of the cold ore produced a temporary cooling of the furnace, but within about 15 min. it would again be up to the desired temperature. The temperature of the furnace and the volumes of air and oxygen supplied to the furnace were read every 15 min. The average volume of air supplied (ordinary air + oxygen) in all but one of the experiments was 30.8 cu. ft. per hr. In the one experiment referred to, for which enriched air containing 42 per cent, oxygen was used, one-half the usual volume was supplied, or 15.4 cu. ft. per hr. Samples of the air supply, when enriched air was being used, were taken occasionally and analyzed for oxygen; the variation in the oxygen content was never more than a fraction of a per cent, during the course of an experiment. Samples of the, roaster gas were taken every half hour and analyzed for SO2. Samples of the ore were taken, usually, at intervals of 1¼ or 1½ hr. These were analyzed for total sulfur and water-soluble sulfur; the latter is approximately equivalent to the sulfur present as normal zinc sulfate.

Data Obtained from the Experiments

In Figs. 9, 10, and 11 are plotted the data obtained from a series of roasts made at the constant temperature of 800° C. This series includes one roast with ordinary air, one with enriched air containing 28 per cent, oxygen, one with enriched air containing 42 per cent, oxygen, in all of which the volume of air supplied was 30.8 cu. ft. per hr., and one roast with enriched air containing 42 per cent, oxygen, in which the volume of air supplied was 15.4 cu. ft. per hr. Fig. 9 shows the variation of the SO2 content of the roaster gas as the roasts progressed; Fig. 10 shows the progressive decrease in total sulfur content of the ore; and Fig. 11 the variation in water-soluble sulfur content of the ore. It should be noted that the vertical scale in Fig. 11 is ten times that in Fig. 10.

Theoretically, if the volume of air supplied is the same, the rate of the oxidation reaction, and consequently the SO2 content of the roaster gas, should vary directly as the partial pressure of oxygen in the air supplied for roasting. When air containing 28 per cent, oxygen is supplied, the SO2 content of the roaster gas should be 33 per cent, greater than when ordinary air containing 21 per cent, oxygen is supplied; and with air con¬taining 42 per cent, oxygen, the SO2 content of the roaster gas should be doubled. The time required for roasting should be in inverse ratio to the partial pressure of oxygen in the air supplied. As shown in the first three columns of Table 5, this is borne out approximately by the experimental data.

By halving the volume of air supplied, keeping its composition the same, the SO2 content of the roaster gas can be increased considerably., but the time required for roasting is also increased by about 50 percent., as will be seen by comparing the last two columns of Table 5, and the curves in Figs. 9 and 10.

In the roasting of this ore, made up of fairly evenly sized particles, the SO2 content of the roaster gas was fairly constant until most of the sulfur was eliminated from the ore, especially when air of moderate oxygen content was supplied. This would probably not hold true when roasting an ore made up of a mixture of fine and coarse particles. The total sulfur content of the ore decreased at a uniform rate in all the experiments, until it was reduced to between 1 and 2 per cent., after which it decreased very slowly; this agrees with the usual experience in roasting in practice. In the roast in which half the usual volume of air was supplied, the sulfur content of the ore, when sulfur elimination stopped, was over twice what it was when the larger volume of air was supplied.

The curves in Fig. 11, showing variation of the water-soluble sulfur content of the ore, are interesting. This sulfur remained fairly constant at between 0.1 and 0.2 per cent, in all experiments until the total sulfur

content of the ore became very low. It then increased sharply to a maximum and later decreased again, first sharply and then more slowly, with continued heating. The height of this maximum, and the amount of water-soluble sulfur remaining in the ore at the end of the roast, increased with increasing oxygen content of the air used for roasting. This agrees with the observation made, as a result of the preliminary laboratory experiments concerning the effect of oxygen on the rate of oxidation of sphalerite. Decreasing the volume of air supplied per hour greatly increased this tendency to form zinc sulfate.

Before running the above experiments at 800° C., some similar roasts were made at 750° C., but in this series the mistake was made of charging the ore in the cold furnace and heating the latter up to roasting temperature afterward. Thus, a variable amount of sulfur was eliminated before the furnace reached 750° C. and, while the results were similar to those obtained in the roasts at 800° C., the separate experiments are not strictly comparable with one another. For that reason the analyses of SO2 in the roaster gas are not given, but the curves showing the rate of sulfur elimination from the ore in the final stages of the roast are of such interest that they are given in Fig. 12. The water-soluble sulfur is here plotted on the same scale as the total sulfur, as it runs considerably higher than in the roasts at 800° C.

Noticing first the curves showing the variation of the water-soluble sulfur content of the ore as the roasts progressed, it will be noted that, as in the roasts at 800° C., the water soluble-sulfur remained very low until most of the sulfide sulfur was eliminated from the ore, and then increased sharply to a maximum that was considerably higher than in the roasts at 800° C. Instead of again decreasing rapidly, as at 800° C., it remained stationary at the maximum or at least decreased only very slowly with continued heating. Thie tendency for zinc sulfate to be formed is greater at 750° C. than at 800° C., and the sulfate is not so readily broken up again at the lower temperature. At this temperature, as at 800° C., the formation of zinc sulfate was greater in the roasts with enriched air of high oxygen content.

The curves show that the total sulfur content of the ore decreased at a uniform rate until it was reduced to a few per cent. Then the rate of sulfur elimination became slower, at the same time that the water-soluble sulfur began to increase. Finally, the total sulfur in the ore actually increased and followed along parallel with the water-soluble sulfur. The explanation of this would seem to be about as follows:

The rabbles used in these earlier experiments, though they kept the ore spread evenly over the hearth and thoroughly mixed, for the most part, left a small amount of ore caked in the corner formed between the floor of the muffle and the circular wall. This ore roasted more slowly than the rest and continued to give off SO2 after the rest of the ore was almost completely roasted. This SO2, together with the oxygen of the air, especially in the roasts with enriched air, reacted with the zinc Oxide in the main portion of the ore to produce zinc sulfate, to such an extent that the total sulfur content of this main portion of the ore increased.


It may be concluded, from the data obtained from these roasting experiments, that temperature, volume and composition of air supply, rate of rabbling; and other such conditions being equal, the rate of oxidation of a given zinc ore increases approximately in direct proportion with the oxygen content of the air supply; consequently that the SO2 content of the roaster gas varies approximately directly, and the time required for roasting varies inversely, as the oxygen content of the air supply. If air containing a high percentage of oxygen is supplied, but in reduced volume, roaster gas very high in SO2 can be produced, but in this case the time required for roasting is considerably greater than when air of the same composition is supplied in the usual volume. In other words, the use of enriched air in roasting can be expected to give a proportionate increase in both SO2 content of the roaster gas and rate of roasting, but an extremely high SO2 content in the roaster gas can only be obtained by sacrificing the gain in the rate of roasting, and vice versa.

The tendency to form zinc sulfate is greater with enriched air than with ordinary air.

Results that may be Expected from Application of Oxygen Enriched Air to Zinc Roasting in Practice

It is in the roasting of zinc ores for the manufacture of sulfuric acid that the use of enriched air would be of particular benefit and, at least as far as we can foresee at present, the possibility of the practical application of enriched air to zinc roasting is not great except where the sulfur dioxide in the gas is to be made use of in some way. In this country, the Hegeler kiln is almost universally used for roasting zinc ores when the roaster gas is to be used for making acid; hence it is chiefly the application of enriched air to roasting in Hegeler kilns that will here be considered.

The possible advantages to be derived from the use of oxygen-enriched air in zinc roasting are an increase in the capacity of the roasting furnace, a decrease in the fuel consumption of the roasting furnace, and an increase in the SO2 content of the roaster gas. From the increased SO2 content of the roaster gas would follow increased capacity and more economical operation of the acid plant.

Our experiments in roasting with enriched air in a laboratory roaster show that with equal temperature, rate of rabbling, and volume of air supplied, the SO2 content of the roaster gas and the rate of roasting increase in the same ratio as the oxygen content of the air supply. The heat balances given in the first section of this paper (Cases 2 and 3) show that increases in the rate of roasting of 95 per cent, when enriched air containing 25 per cent, oxygen is supplied, and 58 per cent, when enriched air containing 50 per cent, oxygen is supplied, are necessary to obviate the necessity of using fuel. In obtaining these figures, the SO2 contents of the roaster gases were assumed as 13.3 per cent, and 28.56 per cent., respectively. Our experiments indicate that the SO2 content of the roaster gases cannot be raised this high except by greatly reducing the volume of air supplied; and if this is done, the capacity of the roasting furnace is correspondingly reduced. It would seem then that large roasting capacity and roaster gas with high SO2 content cannot be simultaneously obtained except by the use of enriched air of very high oxygen content and that, therefore, the use of fuel cannot be done away with except by the use of such highly oxygenated air.

It should be borne in mind, however, that rabbling in a Hegeler kiln is done only at very infrequent intervals and that the ore is therefore very inefficiently exposed to the current of air passing over it. If it could be arranged to use enriched air and rabble, let us say, twice as frequently, the rate of roasting and SO2 content of the roaster gas would be much increased and roasting without the use of fuel would be more nearly within the realm of possibility. This question can only be decided by experiments with a roaster having a capacity approaching that of a full-size furnace.

The possibility of applying enriched air to Wedge furnaces, such as those in which the autogenous roasting of zinc ore is now being attempted, should also be mentioned. Roasting can be carried on autogenously in these furnaces as long as everything goes just so, but the margin of heat is so small that any disturbance of conditions in the furnace is apt to upset the balance. The use of air only slightly enriched in oxygen would increase the margin of safety so that no provision would be necessary for burning fuel in these furnaces.

In conclusion, the writers wish to state that, while they believe that the experimental data and the heat balances which they have given are reasonably accurate, they realize that their interpretation of them is not the only possible one and that from the same data, other metallurgists may draw different conclusions as to the effect that the use of enriched air may have on roasting zinc ores in practice. It is hoped that the data given may be of help to others who are working on the application of oxygen-enriched air to the same or similar phases of metallurgy, and serve to stimulate further thought on the subject.


Copper Geology and Mining Methods

The Chitina mining district of Alaska is located at the headwaters of the Chitina and Copper Rivers. At present, the only producing mining properties are the mines of the Kennecott Copper Corpn. and the Mother Lode Coalition Co., which are situated 196 miles from Cordova the port of entry.

The first claims, later acquired by the Kennecott Mines Co. and afterwards transferred to the Kennecott Copper Corpn., were discovered in 1900. The Copper River & Northwestern Ry., which connects the mines with tide water at Cordova, was completed in the spring of 1911.

Contemporary with the construction of the railroad, aerial tram equipment was brought to the mines by pack train and a tramway, 3 miles long, connecting Bonanza mine with the proposed railroad terminal, was finished, enabling shipments of high-grade ore to be made immediately on the completion of the railroad. A mill to treat the lower grade ore was begun the same year.

The Kennecott company’s holdings consist of 111 mineral claims. The Mother Lode Coalition Mines Co., which is controlled by the Kennecott Copper Corpn., owns 73 claims adjoining the Kennecott holdings. All data on operations and geology refer equally well to the Mother Lode property.


The general geology of the district has been covered by the U. S. Geological Survey and the geological features of the mines have been carefully studied by A. M. Bateman, in his capacity as consulting geologist to the company.

The formations in the vicinity of Kennecott are shown, by the U. S. Geological Survey, to be as follows:

Quaternary.—Alluvium: flood plain gravels, sands and silts.
Rock glaciers: broken rock and ice.
Moraines: glacial till, partly sorted.

Jurassic or later.—Quartz diorite porphyry: stocks, sills, and dikes.

Upper Jurassic.—Kennecott formation: shales, sandstones, and conglomerates.

Upper Triassic.—McCarthy shale: shale with few thin-bedded limestones.
Chitistone limestone: massive limestone mostly magnesian, ore containing.

Triassic-—Nikolai greenstone: altered basaltic lava flows.

The Nikolai greenstone is a succession of altered basaltic lava flows, its total thickness, exposed in the vicinity of the mines, is at least 3500 ft. and the base cannot be seen. Numerous prospects have been opened on copper showings in this formation, the ore being usually bornite, chalcopyrite, and occasionally chalcocite; however they have not resulted in productive mines. Native copper is known to occur in all placer operations in gulches cutting the greenstone, some of the nuggets weigh several hundred pounds. In the vicinity of the mines, the strike of the greenstone is N 60° W and its dip 23° to 30° to the northeast.

Chitistone Limestone.—All the important orebodies are in this formation. It is a conspicuous heavy-bedded formation intersected by numerous systems of fracturing; weathering along these fracture planes produced a very rugged topography. It conformably overlies the Nikolai greenstone and is estimated, by Moffitt, to be about 3000 ft. thick.

The lower part of the formation consists of a 4-7-ft. bed of shale; above the shale is 12 ft. of thin bedded, smooth, hard, gray argillaceous limestone, then 23 ft. of thin-bedded, rough, pebbly limestone, containing flattened, cylindrical, fossil-like grains which, from its appearance, Bateman has termed “crinkley lime,” and 30 ft. or more of dull gray limestone. The remainder of the formation consists of massive beds of sparkling light-gray dolomitic limestone, with occasional beds of darker rock. The upper part of the Chitistone limestone becomes thinner bedded and shaly, gradually grading into the overlying McCarthy shales.

Porphyries.—Light-colored quartz diorite porphyries intrude the greenstone and all the sedimentary rocks in the form of stocks, sills, and dikes. They occur most abundantly about one mile from the Bonanza mine, where they form a larger stock, which constitutes Porphyry Mountain.

Faults and Fractures

There are numerous faults both parallel to and traversing the bedding of the sedimentaries. The former are known as flat faults; the latter also pass into and displace the greenstone. There are many displacements of from 1 to 25 ft., and several faults caused a displacement of as much as 1300 ft. Most of these were pre-mineral; however, in the Bonanza and Mother Lode mines there are several instances where a portion of the ore- body has been displaced. Bateman considers that the flat faults have had a direct bearing on the deposition of the ore, the selvage or gouge contained in them acting as a dam to the orebearing solutions.

Ore Deposits

The general geological features and the relative position of the mines are shown in Fig. 1. The orebodies are typical replacement deposits in the limestone, the outstanding features being the intensity of the mineralization and the fact that chalcocite is the predominating mineral in the deposits. As usual, deposition took place along a fissure, or series of fissures that seemingly start from the greenstone, contact.

These fissures have a strike varying from N 30° E to N 80° E and have no definite dip, varying from nearly vertical to 40° from the vertical, most of them more closely approach the vertical, however. The ore-bodies have the same strike and dip as the fissures, although often when a fault plane is intersected, they widen out along these planes and form what are termed the “flat orebodies,” and are identical with the “Manta” orebodies of the Mexicans. The mineralization along the fissures is much less as the fissure passes into the dull gray limestone, and in only two or three instances is any ore found in this formation or the “crinkley lime” beds that immediately overlie the greenstone.

In the Jumbo mine, a fault roughly following the contact between the dolomitic and the dull gray limestone is the west limit of an orebody, the largest mass of high-grade ore so far encountered. This deposit had a cross-section of 80 by 100 ft. and extended from the 150-ft. to the 700-ft. levels, of which a portion 50 ft. wide and 50 ft. high, extending from the 300-ft. to the 600-ft. level, was practically pure chalcocite.

The lower 1000 ft. of the dolomitic limestone appears to be the most favorable zone for ore deposition. All the productive orebodies lie in it and have their greatest width in the lowest beds, gradually becoming , smaller and of lower grade as they extend east into the upper beds. The eastern extension of the fissure is usually filled with calcite. Thus, the orebodies have a rake or pitch practically paralleling the greenstone contact.

There is every degree of intensity of replacement, from large bodies of practically pure chalcocite and its oxidation products, covellite, azurite, and malachite, to the lime containing small bunches or veinlets of these minerals too low grade to mine. There are no defined walls; the grade of the ore is the limiting factor in mining.

In width, the orebodies vary from a few feet to over 100 ft., not including the local widening of the flat orebodies, which sometimes extend another 100 ft.; in length they vary from 150 to over 1000 ft. In some places, practically the entire width is high-grade ore with only a few feet

of lower grade; in others, the high-grade is in veins from 1 to 10 ft. in width, which are separated from one another by lower grade ore. As the eastern ends of the orebodies are reached, with but one exception, no high-grade deposits are found. There are several places where it would appear that pre-existent fissures or veins were filled, but this occurrence is rare.

The Glacier mine exploits a unique and interesting orebody. It is made up of ice, limestone, some greenstone, and chalcocite. The outcrop of the Bonanza mine was a massive deposit of chalcocite located on the edge of a small amphitheater; the debris, resulting from disintegration of this orebody and country rock, fell into this basin and was occluded in a glacier, which now partly fills it. The orebody is 800 ft. long and 85 ft. wide, and the broken ore in payable quantities extends to a depth of 40 ft.; 45 per cent, of the volume is ice, the remainder is broken country rock and chalcocite with a small amount of carbonate ore.

The principal mineral is chalcocite and its oxidation products covellite, malachite, and azurite. Enargite, bornite, and chalcopyrite are occasionally found together with cuprite, luzonite, and other rarer copper-bearing minerals. During the past five years, the ore produced has averaged 70 per cent, sulfides and 30 per cent, carbonates. The ore is divided in two grades: that which is shipped direct to the smelter and the lower grade ores, which are treated in the mill and leaching plant. The high-grade shipments average between 50 and 55 per cent, copper.

Silver exists in the ore in the ratio of about 1 oz. silver to each 130 lb. copper.

General Description

The Jumbo and Bonanza mines are located on the greenstone-limestone contact at an elevation of 6000 ft.; the Erie mine, on the same contact, is at an elevation of 4500 ft.; and the Mother Lode mine is at an elevation of 5200 ft. This last mine was opened in the higher beds of limestone, the vertical shaft intersecting the contact at an elevation of 4400 ft. Contrary to all expectations, the temperature at the elevation of the mine is not extremely cold, rarely falling below —20° F. and during the winter is often 40° warmer than at the mill camp 4000 ft. lower. Freezing or near freezing temperatures prevail even at the lowest levels of the mines, so the mines are dry and dusty; veins of ice are commonly encountered. The only pumping required is during the summer months, when the snow melts and a small part of the water finds its way through open fissures to the upper levels.

The topography is extremely rough and rugged; snow lies on the ground nine months of the year and snow falls throughout the year. Because of the topography, space for bunk houses and other buildings is limited. All hoists, compressors, and other machinery are located underground. Aerial tramways transport the ore to the mill or railroad terminal, all supplies to mines, and, during the winter months, carry all the passengers to and from the mines.

All the mines, except the Erie, are connected underground; a tunnel is now being driven to connect, this mine with Jumbo. Jumbo and Bonanza mines are opened by inclined shafts paralleling the dip of the greenstone and are located about 50 ft. above the contact. These shafts are 14 ft, wide, have two skipways arid a manway, and are 7 ft. high above the rail. The shaft of the Jumbo mine has, a slope distance of 3051 ft. and the shaft of the Bonanza 2416 ft. On account of the flat dip, the manways have stairways in place of ladders.

The skips used at Jumbo have a capacity of 80 cu. ft. and those at Bonanza, 60 cu. ft., with a track gage of 40 in. in both shafts. The Mother Lode mine was opened by, a two-compartment vertical shaft 800 ft. deep. A new incline shaft has been sunk a slope distance of 1405 ft., after the same manner as at the other mines. All are located underground, being connected with the surface by a tunnel. On account of the flat pitch of the orebodies, the vertical shafts would require an excessive amount of development work to open the various levels.

Formerly, levels were driven each hundred feet, this distance was increased to 200 ft., which was found to be too great, and 150 ft. has been accepted as the best distance, all things considered. Two or three pockets are commonly cut at each level and the skips loaded by chutes without a measuring hopper. One pocket for the mill ore is usually capable of holding about 300 tons; the others, for the high-grade and waste, have a capacity of 50 to 100 tons.

Exploration, Sampling, and Estimating

In common with most deposits in the limestone, it is impossible to foretell or estimate accurately the amount or grade of the ore that a block of ground will produce without an unreasonable amount of development work. Diamond drilling has been used to good advantage for exploring unknown ground; in all over 70,000 ft. of drilling has been done. The usual and more reliable method of exploring has been to drive a drift or crosscut in the dolomitic limestone paralleling the strike of the greenstone, and about 100 to 150 ft. from it; thus any mineral-bearing fissure that is encountered can be followed.

Only occasionally is any sampling done underground. After becoming acquainted with the ore, it is possible to estimate closely the grade of the ore by the amount of glance or carbonates it contains. When the limits of the ore are reached, samples are sometimes taken. It has been found that the sample values are usually considerably higher than the actual recovery obtained in the mill; this is probably due to the friability of the glance and the soft chalky nature of some of the carbonates.

Mining Methods

The shrinkage method of stoping has been used, except for the open-pit mining on the Bonanza mine outcrop. A departure from the usual method, however, is practiced. Where the high-grade portion of the orebody is of sufficient size, as much as possible is mined by the shrinkage method and completely drawn out. The mill-grade ore is then stoped, filling the void left by the extraction of the high-grade and the excess is drawn off as usual.

After as much of the high-grade ore is mined as is practical, other veins, lenses, and masses are met and broken with the mill ore. No attempt is made to sort the ore in the stopes after the mining of the mill ore is commenced; but at all the mines, the ore from the skip pocket on the top level passes over a picking belt, where pieces of high-grade ore are hand picked from the mill ore and any mill ore that may be mixed with the high-grade produce is picked out.

The character of the ground makes almost an ideal condition for the method employed. The work must be given close attention to guard against leaving ore that makes along bedding planes, faults or cross fissures, away from the main orebody; although in most instances as the broken ore is drawn from the stope, it is safe to follow it down and, by using a Jackhamer, recover the ore that may have been overlooked. A great many of the floor pillars left are recovered after a level is finished; but it has been found that it is well not to be too hasty about the recovery of pillars and destroying the level, as oreshoots from a lower level have been found in ground that was considered barren. Until recently, no attempt was made to fill these old stopes, as they would stand empty with practically no caving; the waste from development work is now being used for this purpose.

The Glacier mine is worked but three months per year, when surface mining is carried on. During the months of July, August, and September, the ice of the glacier melts sufficiently to release about 30,000 tons of ore; this is recovered by scraping the thawed ground with a Bagley scraper. To date, while some experimental work has been done, thawing by artificial means has not been attempted; possibly operations might be successfully carried on during the cold months, but it would be at a much greater cost. The scraper used has a capacity of 50 cu. ft. and is operated by an electric double-drum engine of 75 horsepower.

Development Plans

As the inclined shafts are located on the western limits of the ore, crosscuts are driven until the orebodies are reached. The drifts on the ore are kept, as far as possible, in the high-grade ore, chute raises are driven 25 to 35 ft. apart, and widened in the usual manner so that they connect, leaving a pillar 25 to 30 ft. thick between the level and the bottom of the stope. Often, if the ore becomes leaner in the drift, work in the stope is carried ahead from the last chute raise, thus determining the direction in which the drift should be driven. In the wide portions of the orebody, a second, and sometimes a third, drift is necessary to draw the ore evenly from the stopes. In other words, the main idea, after the ore is located on a level, is to follow it, as local swells and pinches in the orebody and the method of mining followed preclude any definite layout of the haulageways as in lower grade and more regular orebodies.

In order to mine the ore on the extreme west end of the orebody, it is necessary to drive raises through the underlying dull gray and crinkley

limestone and the greenstone; when the levels are driven 200 ft. apart, a sublevel is driven to eliminate the long raises that would be necessary.

Fig. 2 shows, in plan and projection, a typical orebody and the development work required to stope it. The main haulageways are driven 7 ft. wide by 7 ft. high on a grade of 0.5 per cent, in favor of the loads; the prospecting drifts and crosscuts are 5 by 7 ft.; 16 and 30-lb. rails are used, the gage of track is 18 in. A compressor plant at Bonanza mine furnishes air for all the connected mines, a 6-in. line being used.

Very little timber is used, only an occasional set being necessary in passing through faults or on the greenstone contact; usually native round timber is used with round poles for lagging.

Loading machines are used in driving the larger headings; while they expedite the removal of the broken material, thus avoiding any delay when the miners are ready to set up for the lifters, a crossbar being used, they have not reduced the cost per ton removed. Scrapers are used at the Glacier mine, as noted; they are also employed advantageously when the main inclines are raised out, instead of being sunk.

Tramming is done by hand, horse, and storage-battery locomotives. Hand tramming is used where the distance is short and a small tonnage is moved; horse tramming, when the distance is greater; for the long hauls and on the levels producing the greatest tonnage, 4-ton Baldwin-Westing-house locomotives with Edison cells are used. This type of locomotive has given very satisfactory service.

For horse and hand tramming, 20-cu. ft. end-dump cars are used; with locomotives, cradle-type and side-dump cars of 36 cu. ft. capacity are used, usually in trains of six or eight cars. .While, on several levels, the locomotives run on 16-lb. rails, the practice is to use 30-lb. rails; curves have a minimum radius of 40 ft.

Hoisting is done in balance, the hoists at the Jumbo and the Bonanza are duplicates; they are of single-reduction, herringbone-gear type with a rope speed of 600 ft. per min., driven by two 85-hp., a.c., 2200-volt, three-phase, sixty-cycle motors; they were manufactured by the Allis Chalmers Co. The Mother Lode incline will be equipped with a double-drum hoist, with double reduction gears, driven by two 75-hp. motors; the rope speed will be 450 ft. per min. The cables are six-strand, nine- teen-wire, Lang lay, 7/8 in., in diameter. When hoisting men, the skips are removed and a man car used. Neither skip nor man car is fitted with a safety device, as a satisfactory one has not yet come to the company’s attention.

The air-compressor plant furnishes air for all mines, except the Erie, where an Ingersoll-Rand Imperial type 10, 600-cu. ft. capacity, electrically driven compressor is installed. The plant contains: One Ingersoll Rand type P. E.-2 compressor, 1500 cu. ft. capacity, driven by a 250-hp. synchronous motor; two Ingersoll Rand Imperial type 10 compressor, 500 cu. ft. capacity, each driven by a 85-hp. motor; one Ingersoll Rand Imperial type 10 compressor, 650 cu. ft. capacity, driven by a 105-hp. motor.

Because of the numerous openings to the surface, natural ventilation, with the exception of small fans belt-driven by a 10-hp. motor in development, aided by doors to course the air, is satisfactory.

Electric lights are used on the levels and incline shafts. The miners use carbide lamps, furnishing their own caps and lamps, the company keeping them in repair.

Each level has a telephone connecting with the foreman’s office, compressor room, and hoist room. The mine telephone system is independent of the general system.

Electric pull bells, modeled after those commonly used in other mines, are used.

Operating Data

Types of Drills

For drifting Ingersoll Rand, 248 Leyner machines are used; for stoping and raising, Ingersoll Rand C. C. 11, except when drilling in chalcocite, when it is necessary to use a water-type drill. Ingersoll Rand B. C. R. 430 and Sullivan D. P. 33 are used for blockholing and where occasional flat or down holes are to be drilled. Four-point, cross, high-center drill bits are used on all machines, made up of the following sizes of steel: 1-in. quarter octagon for stoper; 7/8-in. hollow hexagon for Jackhamer; 1¼-in. hollow round for Leyner. The bits are:

Stoper, 1 7/8-in. for starters; 1 ¾-in. for seconds; 1 5/8-in. for thirds; and 1½-in. for fourths.

Leyner, 2-in. for starters; 1 7/8-in. for seconds; 1¾-in. for thirds; 1 5/8- in. for fourths.

Record of Unit Production

(a) Ore broken……………………………………………………297,502 short tons
Ore produced…………………………………………………….294,202 short tons

Labor Data

(c) Stoping labor includes: Miners in stopes, muckers in stopes, bulldozers in stopes, rockbreakers in stopes:
Tons broken per man per hour…………………………….1.3964
Man-hours per ton……………………………………………………0.7161
(d) and (e) Exploration and development labor, miners only:
Tons broken per man per hour……………………………1.2941
Man-hours per ton…………………………………………………0.7726
(g) All underground labor including above labor:
Tons produced per man per hour……………………..0.4586
Man-hours per ton…………………………………………………2.1807
(h) Surface labor, exclusive of office force:
Tons produced per man per hour……………………11.5536
Man-hours per ton……………………………………………….0.0866
(i) All labor including office force:
Tons produced per man per hour…………………….0.4243
Man-hours per ton……………………………………………….2.3570

Supplies Data

Safety Measures

Hoisting ropes are thoroughly inspected once each week; every six weeks 2 ft. are cut off from both ends. The ropes are changed end for end after six months’ use. The sheaves are inspected once every week, and the hoists each day. In the vertical shaft, the safety catches are tested every Sunday.

There is a fire extinguisher on every level station; fire doors are provided. When located near timber or snow sheds they are of concrete and steel; otherwise they are built of wood, care being taken to make them as air-tight as possible.

No safety engineer is employed, the engineering department reporting to the general mine foreman and superintendent any unsafe practices that come to its notice.

Each bunk house is equipped with a pool and reading room, a number of magazines and other periodicals being provided. Moving picture shows are given twice a week at Jumbo and Bonanza camps.

A well-equipped hospital is located at the mill camp with a competent surgeon and corps of nurses in attendance.

At the plant, last year, there was one fatal accident; no serious accidents causing total permanent disability; three partial permanent disability; 28 causing loss of more than 14 days time; and 120 minor, loss from 0 to 14 days.

Compensation paid, under Territorial Act, amounted to 1.058 per cent, of the payroll.