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.

Conclusions

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.

Conclusions

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.

Conclusions

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.

Geology

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.

Mining Methods and Costs

The Globe Mining District is in the southeast central part of Arizona, in Gila County. Globe, with a population of about 7000, is the terminus of the Arizona Eastern R.R., a branch line 130 miles long that connects with the Southern Pacific R.R. at Bowie.

In 1874, prospectors crossing the Pinal Mountains from the west located what is generally known as the Old Dominion mine. For several years, it attracted little attention, because of the greater interest aroused by the discovery of high-grade silver ores in some of the foothills northeast of Globe. About six years later, the prospector turned his attention to the abundant copper ore revealed by surface workings along the Old Dominion vein, and, in 1884, the Old Dominion company erected two 30-ton furnaces. From 1888 to 1893, the Old Dominion company is said to have maintained an average annual production of about 8,000,000 lb. of copper. Until Dec. 1, 1898, all supplies had to be freighted into Globe by wagons and the mines of the district operated intermittently because of high expenses, but with the advent of the railroad the Old Dominion company continued to be a large and steady producer.

In the Globe district, the production of copper far exceeds in importance that of any other metal. There are four operating companies along the Old Dominion vein and the total annual product of these properties for 1923 was about 45,000,000 lb. of copper.

The unit size of the mineral tracts in the district is the regulation mining claim, 600 ft. wide by 1500 ft. long, and all ownerships are held in fee.

Globe is 3600 ft. above sea level, and lies between the Apache Mountains to the east and the Pinal Mountains to the west. The principal drainage of the district is northward through Pinal creek into the Salt River. The general slope from the high point along the vein, where the Superior & Boston mine is, to Pinal creek, where the Old Dominion mines are, is about 250 ft. to the mile.

Geology

The oldest rocks in the district are pre-Cambrian crystalline schists known as the Pinal schist, which are the basement upon which all the later rocks were deposited. These latter rocks comprise shale, conglomerates and quartzites with a total thickness varying from 500 to 800 ft. and are thought to be Cambrian in age. Overlying these rocks is a series of limestones, known as Globe limestones, that vary in thickness from 300 to 500 ft. and range in age from Devonian to Pennsylvanian.

These rocks have been cut by numerous faults, and following or accompanying the faulting large sills and masses of diabase were intruded between the sedimentary beds. A long period of erosion followed, during which the region was deformed by further faulting to its present topography and during which the original ores were deposited.

The main fault in this district and the one along which most of the mining is carried on is known as the Old Dominion fault; it varies from 3 to 50 ft. in width and is developed for a length of approximately 3 miles. The fissure has a variety of strike and dip but is roughly north-east and southwest, with a dip of about 80° to the south.

The fault is fairly conspicuous and is easily followed, except where it is wholly in diabase, when its course is marked by a zone of brecciation stained with hematite and salts of copper.

The vein is commonly made up of brecciated shale or quartzite and mineralized with oxide of iron and the ores of copper, the overburden varying from 200 to 600 ft. The mineralogical character of the ores along the vein is simple. The oxidation of the sulfides has resulted in simple products. The pyrite and chalcopyrite have their sulfur replaced by oxygen, carbon dioxide, or silica and become hematite, limonite, cuprite, malachite, or chrysocolla. The secondary sulfides recognized in the district are chalcocite and bornite. Native gold, silver, and copper have been observed in small amounts within the zone of oxidation.

Exploration

Most of the exploration along the Old Dominion vein has been done by test pitting, tunneling, trenching, shaft-sinking, drifting, and cross-cutting. All development is carefully sampled, the outline and extent of ore body carefully determined, and ore estimated on a basis of 11 cu. ft. per ton. The production, as indicated by exploitation, has proved the method sufficiently accurate.

Change in Mning Method

The principal mining method formerly in use was the square-set method but the decreasing copper content and the increasing cost of timber, together with the increasing cost of labor and supplies, made it imperative that a cheaper method be substituted. In some places along the vein where the ground is heavy and where it is imperative to keep timber close to the working face, square-setting is used, but in general that method has been superseded by newer and more economical methods; the selection of method depends on the size and shape of the orebody and the character of vein filling and walls.

Sampling and Estimating Reserves

The method of sampling is far from elaborate; grab samples only are taken from each round blasted daily. A record is kept of all samples taken in the block of ore lying between any two raises and the numerical average for the month is assumed to represent the value of the ore mined that month.

When estimating the reserve, which is done the first of every year, the area of each block of ground remaining between any two raises is measured on the profile tracings with a planimeter. This area multiplied by its average width gives the contents in cubic feet. This figure divided by 11 (11 cu. ft. ore in place is equivalent to 1 ton) is reported as the tonnage for that block. The numerical average of the year’s samples of ore broken, together with the assay values of the drift over this section, is reported as the assay value of the block remaining.

The total ore reserve is computed by multiplying the number of tons by the per cent, for each block. This product divided by the total reserve tonnage gives the average per cent., which practically checks with the heads reported by the mill. Blocks of ground that have been worked out have been found to check within 0.5 per cent, on both tonnage and value reported.

Development

The section of the Old Dominion vein along which the Iron Cap Copper Co. is mining is about 3500 ft. long, and the width of the vein varies from 3 to 40 ft. with an average dip of about 80°. The vein material is a hard brecciated quartzite or shale between fairly good walls. The distribution of values through the deposit is irregular and some sorting is resorted to. Very few waste bodies occur in the ore zone, however, and when found are usually left in place until stopes are ready for waste filling; they are then blasted down and become part of the gob.

The inclined cut-and-fill system is used throughout the mine. This method requires very little timber; stope floors are carried on an incline of about 34°, which eliminates most of the labor of shoveling but which is not so steep as to constitute a hazard from rolling boulders.

The average stope temperature is about 78°, average relative humidity 88 per cent. Production is approximately 150 tons per 8-hr. shift of 50 men. This includes all underground labor, but only about 50 per cent, of the shift are on actual stoping operations.

Fig. 1 shows the successive steps from the starting of a stope to its finish. This plan calls for a main shaft for the handling of all men and materials and the opening up of the mine by a series of levels placed approximately 100 ft. apart.

The Iron Cap Copper Co. has two three-compartment shafts, each compartment 4½ by 5 ft. in the clear, timbered with 10 X 10-in. timber sets on 5-ft. centers and lagged with 2 by 12-in. lagging. The Iron Cap shaft, the only one at present operating is 1540 ft. deep, and is in the hanging wall. Commencing at the 800-ft. level, stations 15 ft. high by 40 to 60 ft. long are cut every 100. ft. Station sets are 10 by 10-in.

timber on 5-ft. centers with a drop of 6-in. on each set, leaving the back end of the station 11 ft. high.

Loading pockets were cut above the 1100-ft. level and under the 1300-ft. level; raises driven from each pocket accommodate the ore mined on three levels.

The shaft is so conveniently located with respect to the vein that all ground between shaft and vein constitutes the length of the station. Drifts 5 by 7 ft. are then driven along the footwall, no timber being used until stopes are started. Fig. 2 shows the arrangement and dimensions of shaft, station crosscut, and drift.

Mining Methods

As soon as the drifts have advanced far enough, raises with a minimum cross-section of 5 by 10 ft. are driven on 125-ft. centers. The sill floor sets of these raises are timbered with 10 by 10-in. timber. Posts are 8½ ft. high, sets placed on 5-ft. centers. Above the sill floor set, however, the timber is 8 by 8 in. Posts are 5 ft. 4 in. long.

The manway only is timbered with what is termed a “clap-me-down” set. The posts are set as nearly over each other as possible,

the cap is cut to fit the ground and is well blocked; 3-in. by 12-in. by 6-ft. lining boards keep the manway clean. A 6 by 6-in. sprag in the chute end flush with the last cap serves as staging for the machine men.

All headings are given definite numbers, which indicate to a certain extent their location with respect to the shaft. Headings to the east of the shaft have an even number and headings to the west, an odd number. Raises are numbered consecutively from the shaft in each direction. For instance, 902 raise No. 4 would be the fourth raise east of the shaft on the 900-ft. level. Stopes are designated as 902 stope east or west of raise No. 4.

Stopes may be started as soon as the raise has been holed through to the level above. The back of the original drift is first broken down to a height of 15 ft. and the ore mined from footwall to hanging wall. This sill-floor stope is then timbered with 10 by 10-in. sets with square framing. Sets are placed on 5-ft. centers; posts are 8½ ft. long. If the vein is not over 8 ft. wide, the cap is cut to fit the ground. In some cases where the vein is 6 ft. wide or less, and walls are exceptionally hard, hitches are cut to receive the cap and no posts are used.

A temporary chute is placed 25 ft. on each side of the original raise; permanent chutes are placed 50 ft. on each side of the original raise with a manway between.

Filling Stopes

As soon as all sill-floor timber has been placed, the sets are lagged over with a double floor of 3 by 12-in. planks and stoping starts on the first floor at the original raise. The ground is blasted out around the raise as high as safety will permit, the ore is then removed and filling poured in from the level above, the waste taking its own angle of repose. When the waste filling is about 3 ft. from the back of the stope it is roughly leveled off and floored with 3-in. by 12-in. by 5-ft. planks. Cleats of scrap timber are nailed to the floor to enable machinemen to move around easily and a cut about 6 ft. high is taken each side of the raise. When this cut is completed, the floor is taken up and piled out of the way and the opening is filled as before; the flooring is again laid and another cut taken. The flooring is used until it is worn out.

As the stope passes the temporary chute, the timber from this chute is salvaged for use elsewhere. As the toe of the incline reaches the permanent chute, the original raise timbers are salvaged as the stope progresses and are used to build up the permanent chutes and manway.

In all main drifts 2-in. air lines and ¾-in. water lines are carried, 1-in. air lines and ½-in. water lines are run down each original raise and up each center manway, so that drilling connections may be made at either the top or the bottom of the incline.

As stopes hole through to the level above, drift timbers are caught up and held in place until they can be supported by 8 by 8-in. sets placed upon the filling; these are eventually filled in and the original drift left intact.

Waste fill for stopes is obtained from development work, from shrinkage stopes above the ore zone, and from old filled stopes where no damage can be done by allowing them to cave.

Drilling and Blasting

All stoping is done with wet hand-rotated stopers using 7/8-in. quarter-octagon hollow steel. The starter bit is 1¾ in. and decreases 1/8 in. on each length of steel. Thirty-five per cent, gelatin powder is used in blasting all holes in stopes.

A grab sample is taken from all rounds blasted and a copy of the assays is furnished to bosses daily. All broken material is hand trammed in 16-cu. ft. end-dump, roller-bearing cars run on 12-Ib. rails.

The method adopted for drilling drifts, raises, and shaft sinking do not call for any special mention. This work is usually done on contract, the company providing all tools, equipment, and supplies; and the contractor providing all labor. The average price paid for drifting is $4.40 per linear foot with a minimum cross-section of 5 by 7 ft. The price for raises with a minimum cross-section of 5 by 10 ft. is $4.50 for the first 50 ft. and $5 per foot for the remainder. Shaft sinking averages about $50 per foot, depending largely on the character of rock being drilled and the amount of water likely to be encountered. All drifting and shaft sinking are done with water Leyners using 1¼-in. hollow round steel with double-taper cross bits. The starter bit is 2 in. and decreases 1/8-in. on each length of steel.

Blasting in all development work is done with 40 per cent, 1 1/8-in. gelatin powder. All development is carefully sampled and accurate assay maps brought up to date every 30 days. About 1 ft. of development is done for every 12 tons of ore mined.

Costs

Tables 1 and 2 show average detailed costs of stoping and development work per ton of ore for 85,211 tons mined in 1923. These costs constitute approximately 50 per cent, of the total cost of mining. At

present 30 miners break all ground in development work and stopes, including waste to fill stopes, and supply 300 tons of ore daily. There is an average of 100 men employed underground and a total of 135 at the mine. All labor in stopes is on a “day’s pay” basis.

Machinery and Surface Plant

The surface equipment consists of an Allis Chalmers double-drum hoist driven by a 250-hp., 440-volt, 25-cycle electric motor, each drum holding 1800 ft, of 1 1/8-in. 6 X 19 Lang lay cable. All hoisting is done in counterbalance at a speed of 700 ft. per min. Ore is hoisted from the pocket through two compartments in 4-ton skips and dumped direct into a choke feed Austin No. 7½ gyratory crusher, which breaks to about 2 in. This material then passes through a trommel and the oversize is fed to a Symons 48-in. vertical disk crusher, the final product being ½ in. A belt conveyor carries it to the ore bins, and from there it is taken to the mill, ½ mile away, by a Westinghouse 6-ton locomotive operating over a 24-in. gage electric railroad on 550-volt d.c. current.

A cage for hoisting men is hung under the skip, and provision is made for connecting a second cage underneath if necessary. An unbalanced dinkey cage is operated in the third compartment by a 10 by 18-in. duplex, direct-acting sing!e-reel Ottumwa hoist; this cage is used only to lower supplies. Skips and cages are both equipped with safety catches and are inspected daily.

Pumping

At present, the water is handled by electrically driven pumps in two lifts, from the 1500 to the 1300-ft. level, and from this level to the mill on the surface. Both pumps are on the 1300-ft. level. A Lane & Bowler 500-gal. deep-well pump, six-stage, driven by a 40-hp. motor lifts the water from the 1500-ft. level through an 8-in. pipe and discharges into two 60,000-gal. concrete sumps on the 1300-ft. level. The water gravitates into a 300-gal. Aldrich quintuplex plunger pump driven by a 150-hp. motor and pumps through a 6-in. pipe direct to the mill on the surface, about ½ mile from the collar of the shaft.

Air Compression

Air for drilling is furnished by a steam-driven 3000-cu. ft. O. R. C. Ingersoll-Rand compressor. A Sullivan 1500-cu. ft. tandem, compound, direct-connected, steam-driven compressor is idle at present but can be used in an emergency.

Ventilation

Ventilation is provided by one Sturtevant multivane fan pulling 65,000 cu. ft. at 3½-in. water-gage pressure, driven by a 75-hp. motor, belt-connected. This exhaust fan is at the collar of the Williams shaft and is operated 14 hr. per day. Air is drawn down through the Iron Cap shaft to the lowest mine level and allowed to work upward through the stopes, finally finding its way out through the Williams workings and up the Williams shaft.

Lighting and Signaling

Stations are lighted by 32-c.p., 100-watt, 110-volt electric bulbs; main tramming drifts are lighted by 16-c.p. 40-watt, 110-volt electric bulbs. Lights in stopes are provided by carbide lamps carried by each miner.

Western Electric, local-battery system telephones are located on each shaft station and at the collar of the shaft. All signals between cagers and hoisting engineer are over an electric signal system, supplemented by rope bell.

All electric power used is bought from the Inspiration Copper Co. at Miami, Ariz., and brought over a private line a distance of 7 mi. All steam power is furnished by three 250-hp. Babcock & Wilcox water- tube boilers.

Discussion

A Member.—What do they put on the filling?

A. L. Walker.—They put lagging and use that lagging until it is worn out, gradually moving it from one place to another.

A. Neustaedter, Roselle Park, N. J.—Do they put up square-set raises?

A. L. Walker.—They put up square-set raises, but use square sets only when it is imperative. If the ground is at all soft or dangerous they use stulling.

A. Neustaedter.—Cribbing would not do.

A. L. Walker.—It might in certain cases. Of course, in the old days, the square-set system was used altogether. In the Old Dominion mine, where the orebodies were 40 or 60 ft. wide and sometimes 60 ft. high, square sets were used altogether.

A. Neustaedter.—Do they work the square-set stopes on the rail too?

A. L. Walker.—Yes; all the orebodies above the eighth level in the Old Dominion were worked with the square-set system. About 1893, however, that system became so expensive that we developed a system of heavy stulling to support the roof whenever possible.