Until recently little had been published on methods of treating any of the uranium ores for the extraction of radium, although a number of patents have been issued both in this country and abroad. Therefore persons interested in methods of extracting radium have had little to guide them, as details of the methods in use have been closely guarded. The effect has been rather to discourage those who might have been inclined to invest money in the production and sale of radium. Consequently, there is a real need of the presentation of all of the necessary details in the successful treatment of uranium ores, and the refining of the finished products.

The different methods heretofore used may be classified under three general heads, as follows: (1) Use of an acid leach; (2) use of an alkaline leach followed by an acid leach; (3) fusing the ore with some material that will break up the ore and make the extraction of the valuable contents possible. These different methods have been discussed at some length in Bulletin 70, but are described here in more detail.

Using Acid Leach

In the method involving the use of an acid leach, the two acids employed up to the present are sulphuric and hydrochloric.

Leaching With Sulphuric Acid

A number of patents have been issued in connection with the use of sulphuric acid. The Fleck, Haldane, and White patent claims that radium, uranium, and vanadium can be extracted successfully in the following way:

Method of Fleck, Haldane, and White

The ore is crushed, preferably to 20 to 40 mesh, by any suitable means, and is agitated with hot sulphuric acid of 15 to 20 per cent concentration, the proportion of acid used depending upon the quality of the ore. As a rule, 400 pounds of 65° B. sulphuric acid diluted to 15 to 20 per cent is sufficient for the treatment of 1 ton of ore. The resulting acid solution contains the uranium, vanadium, and copper, if copper is present, and is preferably filtered or otherwise clarified. The resulting clear, acid solution is then brought into contact with fresh ore, heated, and agitated, the solution being thereby neutralized. At the same time, a part of the uranium, vanadium, and iron is precipitated on the ore as basic sulphates or carbonates, the effect of this precipitation being to enrich the ore, which may be of low grade. If necessary, the neutral solution is again clarified, and constitutes a part of the stock solution suitable for further treatment. The enriched ore that has served for the neutralization of the acid solution, either alone or mixed with fresh ore, is treated with sulphuric acid, as described above, yielding an acid solution, which, after neutralization, is added to the stock solution.

The ore residues from the treatment with sulphuric acid, as well as the residues from the similar treatment of the enriched ore, are freed from remaining values by being washed with dilute sulphuric acid. The resulting acid washings are then strengthened by the addition of sulphuric acid until they contain, preferably 15 to 20 per cent of the acid, and are utilized for continuing the process.

The nearly neutral stock solution, containing uranium, vanadium, iron, and possibly copper, is then treated with sulphurous acid by subjecting the solution to the action of sulphur dioxide obtained by roasting sulphur or sulphide ores. This treatment reduces the iron and the vanadium. Reduction of the iron to the ferrous state is an advantage, because in the subsequent precipitation of the uranium and vanadium less iron is precipitated, and the valuable metals are therefore in a more concentrated form. The sulphuric acid from the sulphur dioxide is also available for the treatment of additional quantities of ore.

The reduced solution is separated from the ore by filtration or decantation, and is treated with such a quantity of finely pulverized limestone as calculation has shown will bring it to the point at which uranium, vanadium, and copper will just begin to be precipitated, calcium sulphate being formed. The solution is then separated from the calcium sulphate, and sufficient pulverized limestone is added to cause complete precipitation of the uranium and vanadium when the liquid is boiled. The precipitate, which comprises a complex mixture containing basic sulphates and carbonates of uranium and vanadium, compounds of iron, and hydrated calcium sulphate, is at first green, but in air changes rapidly to light green or yellow. This precipitate may be used as the final uranium-bearing product, or it may be further refined or concentrated by any known or suitable method, as the radium, of course, remains with the insoluble residue.

The residue may be agitated with water, and after the larger part of the coarse material has settled, the slimes may be separated and allowed to settle. After the decantation of the clear liquid, the residue, consisting largely of calcium sulphate, barium sulphate, and finely divided silica, constitutes the radium concentrate. From this concentrate the radium may be extracted and refined by any suitable method. By the procedure outlined, it is possible to get a radium concentrate carrying 50 to 100 milligrams of radium per ton of concentrate, the exact proportion depending on the grade of the ore treated.

PART 1: Radium, Uranium & Vanadium Extraction & Recovery from Carnotite

PART 2: URANIUM, RADIUM & VANADIUM Ore Processing

PART 3: VANADIUM & URANIUM Extraction and Recovery

PART 4: RADIUM Extraction & Recovery

PART 5: Processing, Extraction & Recovery of RADIUM 

Methods

The other methods of leaching with sulphuric acid are more or less similar to that of Fleck, Haldane, and White. Some involve a preliminary roasting of the ore before treatment with acid.

McCoy’s method involves mixing the ore with either dilute or concentrated sulphuric acid to form a mud, and roasting, at first to a temperature of 100° C. and finally to 300° C. or even higher. When the roasted material is leached with water, the iron, vanadium, and uranium are dissolved, and the radium remains with the insoluble residue from which it can be recovered by any suitable means.

In all such methods of leaching the radium remains with the insoluble residue, and usually some sliming method is used to obtain it in a more concentrated form.

If carnotite is treated with hot concentrated sulphuric acid, the radium, instead of being precipitated, is dissolved. If the solution obtained from this treatment is diluted and allowed to stand, radium and barium are precipitated, more barium chloride being added, if necessary. In order to get a good extraction it is necessary to use two to three times as much acid as ore.

Theoretically, it might be possible to get radium into solution by passing carbon dioxide into water in which carnotite ore is suspended. This method, however, does not give satisfactory results.

Leaching with Hydrochloric acid

Hydrochloric acid has been used for leaching in a number of commercial plants. The ore is boiled with hydrochloric acid—at some plants with the strong acid, and at others with weak acid—one part of strong acid to two parts of water. As compared with sulphuric acid, hydrochloric has the advantage of putting the radium into solution at the start. Also, as hydrochloric acid has a reducing action on the vanadium, it acts as a good solvent for that metal as well as for the uranium. If strong acid is used, the radium can be precipitated from the acid solution by partial neutralization either with sodium hydroxide or sodium carbonate, preferably the former, and the addition of sodium sulphate and barium chloride if the ore does not contain sufficient barium. After the separation of the precipitated radium-barium sulphate, the uranium and vanadium can be recovered by any suitable process.

Use of Alkaline Leach

It is probable that some of the early experimental work on extracting radium from carnotite ore involved the boiling of the ore with sodium carbonate, thereby getting rid of most of the uranium and part of the vanadium. The residue, after having been thoroughly washed, was then leached with dilute chemically pure hydrochloric acid in order to extract the carbonates present, of which radium would be one. The Haynes-Engle process for the recovery of uranium and vanadium covered this method, but the patent specifications do not show what is necessary for the final recovery of the radium, as at the time the patent was issued carnotite was mainly valuable for uranium and vanadium and not for radium.

Haynes-Engle Process

According to the Haynes-Engle patent the ore is first crushed to 12 mesh and is then boiled with a solution of an alkaline carbonate, preferably sodium carbonate or potassium carbonate, until the uranium or vanadium, or both, in the ore is dissolved. The strength of the sodium carbonate solution and the length of time necessary to boil it are determined by the proportion of uranium and vanadium in the ore, and will probably vary considerably. The originators of the process claim, however, that 100 pounds of sodium carbonate per ton of ore for each 1 per cent of uranium and vanadium, or either, present will give good results. The time required for boiling should be about one hour. After the uranium and vanadium, or either, has been dissolved, the clear solution is drained into a separate tank. The uranium is precipitated as sodium uranate by the addition of sodium hydroxide to the solution. This precipitate is removed from the solution, which contains all of the vanadium. From the solution, either with or without neutralization, the vanadium is precipitated as calcium vanadate by the addition of water-slaked lime.

It is claimed that an extraction of 80 per cent of the uranium and 60 to 65 per cent of the vanadium can be obtained by this process.

Bleecker’s Process

What was lacking in the Haynes-Engle process has been covered by a patent issued to Warren F. Bleecker. After the ore has been boiled with sodium carbonate, as in the Haynes-Engle process, and washed, the residue is leached with a dilute acid, preferably hydrochloric or nitric. This treatment of the ore dissolves the radium, forming radium-barium chloride, and the solution can be drawn off and stored for any approved treatment to recover the radium. If desired, the ore may be leached once more with acid, 10 per cent sulphuric acid being used, the object being to recover the vanadium not extracted by the first leaching.

By Bleecker’s method not only is nearly all of the uranium and part of the vanadium recovered but the radium is also obtained in soluble form. As the bulk of the liquid would be too great for direct fractionation, it would be necessary to precipitate the radium and barium salts in solution by adding a sufficient quantity of sulphuric acid or sodium sulphate. By this means, practically all of the radium in solution could be recovered as radium-barium sulphate, which would need further treatment, as described in a subsequent chapter.

Another patent a issued to Bleecker covered some modifications of the process outlined. Under the later patent the ore is boiled with a solution of an alkaline hydroxide, and, if desired, an alkaline carbonate may be added to the solution. For example, sodium hydroxide and sodium carbonate may be used, in which event the resulting liquid will be a solution of sodium vanadate, from which the vanadium may be recovered by any approved process. After the sodium hydroxide in the ore has been neutralized with acid, the ore is leached with a solution of alkaline carbonate, the temperature of the solution being kept at about 90° C. This treatment will dissolve the uranium as the soluble sodium uranyl carbonate. The solution can be subsequently treated by any suitable process for the recovery of the uranium. After the ore has been thoroughly washed to free it from soluble carbonates and sulphates, it is leached with an acid other than sulphuric, then washed with water. Eight per cent hydrochloric or nitric acid gives good results. By this method the radium is extracted as the soluble chloride in conjunction with barium, calcium, etc., and can be precipitated from the solution, as already described, by the addition of sulphuric acid or sodium sulphate.

Extraction of Radium by Fusion Methods

Fusion With Sodium Sulphate

The first method that was used in connection with the extraction of radium was that of fusing pitchblende ores with sodium sulphate. This method was originally used by the Austrian Government in connection with the treatment of the pitchblende ores from the Government mines at St. Joachimsthal. By this fusion the uranium in the ore is changed to sodium uranate, which can be dissolved from the insoluble residue, after leaching with water, by means of dilute sulphuric acid. The radium remains behind with the residue, and was, before the discovery of radium, discarded.

Procedure of Haitinger and Ulrich

The extraction of the radium from this residue is well described by Haitinger and Ulrich, and, with some minor changes, is probably similar to the method now used by the Austrian Government. The work was done in the laboratory of the Austrian Incandescent Gas-light & Electric Co. Ten thousand kilograms (22,000 pounds) of pitchblende residues was treated, representing about 30,000 kilograms (66,000 pounds) of pitchblende containing 53.4 per cent of U3 O8. Moisture in the material in the various shipments varied from 10.3 to 18.4 per cent. The moisture was driven off at a temperature of 105° C., the 10,000 kilograms of residue losing 1,340 kilograms in this way. The work took two years, owing to time spent on analyses at the start and to experiments to obtain the best results. The method given was developed as being most suitable to the equipment available. Five thousand kilograms annually could be treated.

The chemical operations involved were, first, the decomposition of the residues; second, the removal of the radium sulphate by precipitation; and third, the solution and concentration of the radium sulphate. The first step consisted of digesting the residue with sodium hydroxide, 100 kilograms being boiled for one day with a solution of 50 kilograms of hydroxide in 200 liters of water. Forty per cent of the alkali was converted to sulphate and to silicate. The solution contained some radium, but the total radium so dissolved from the entire 10,000 kilograms of residue represented the radium equivalent of only 10 kilograms. Therefore this solution was thrown away.

After the boiling the residue was allowed to settle and the liquid was decanted. The residue was washed to remove the greater part of the sulphates, filtering and washing being done in a funnel with a capacity of 100 kilograms and having a lead suction tube 3 meters long. The vessel containing the washed residue was placed on a water bath, and the residue was treated with an equal weight of dilute (1:1) crude hydrochloric acid. After prolonged heating the acid solution was decanted and the residue washed with water. The water was then used to dilute the next part of acid for a new sample.

Crystals of calcium sulphate and lead chloride formed in the acid solution as it cooled. Neither the solution nor the crystals contained an appreciable quantity of radium, but nearly all of the polonium and of the actinium was in the solution. The solution was therefore treated with ammonium hydroxide to precipitate the polonium and actinium. The filtrate was not radioactive and was discarded.

The residue from the treatment with crude hydrochloric acid was boiled in a solution of sodium carbonate, the carbonate, made by the ammonia process, being free from sulphates. Fifty kilograms of sodium carbonate in 200 liters of water was used for a 100-kilogram sample. By this treatment a large part of the radium sulphate was converted to radium carbonate. Therefore, in subsequent treatments the solutions had to be kept free from sulphate. The residue was washed till free from all trace of sulphate and then treated with pure hydrochloric acid. The boiling with soda and the treatment with acid was repeated three times. After the third treatment only 2 per cent of the original radium content remained in the residue, which was thrown away. The soda extracts were practically free from radium. The washing of each of the various residues consumed four to six weeks. The hydrochloric acid extracts, containing nearly all the radium, were united and the radium was precipitated as a sulphate with sulphuric acid. Besides the radium, the precipitate contained, of course, the alkali earths, including calcium, much lead containing radio-lead, and a small quantity of rare earths containing actinium. The sulphate precipitate, called crude sulphate, represented 0.5 to 2 per cent of the weight of the original residue.

The crude sulphate was reconverted to carbonate by repeated boiling with sodium carbonate solution. All of the sulphate from any one sample could not be converted; therefore, the final residue from the treatments was not thrown away, but added to a new part of crude sulphate. After each carbonate treatment, an extraction with hydrochloric acid was made. The lead chloride formed in the solutions was removed and freed from radium by repeated crystallization in hot water. Sixty kilograms of lead chloride was thus obtained from the entire 10,000 kilograms of residue. It was saved and treated for its content of radio-lead.

The hydrochloric, acid solutions from the crude sulphate were freed completely from lead by hydrogen sulphide and were then evaporated to dryness on a steam bath. The calcium chloride in the residue so obtained was dissolved in concentrated hydrochloric acid, in which barium chloride is only slightly soluble and radium chloride is still less soluble. The residue remaining, called crude chloride, consisted of radium and barium chlorides, with some strontium and calcium chlorides and traces of other impurities.

From this point on the concentration was continued by fractional crystallizations from water solutions. Radium chloride, which is the least soluble of the chlorides, accumulated in the crystals, the foreign matter remaining more and more in the mother liquor. The first fraction was, of course, the richest in radium. Two steps in this process were of particular importance—first, the separation of as large a quantity of radium-free barium chloride as was possible; second, the making of a relatively large first fraction. The second step could be satisfactorily attained by temporarily stopping the crystallization of the first series until the crystals of the second series were of sufficient radioactivity to be united with the first.

The crystallizations were all carried out on a steam bath in order to avoid contamination with sulphate, as might have occurred if heating had been done with a direct flame.

Finally, two portions of crystals were obtained, one of about 2 kilograms, containing nearly all of the radium, and the other of about 11 kilograms, containing little radium.

The 2-kilogram portion was treated as raw material for the production of radium chloride free from barium. After this portion had been crystallized about 30 times, the first fraction of about 9 grams was further crystallized, and the lower fractions were combined into three groups according to their activity. The 9-gram portion was first purified with hydrogen sulphide, which removed traces of lead that probably came from the glassware. Further work was conducted in quartz vessels. The salt was dissolved in dilute hydrochloric acid, warmed, and allowed to crystallize. Four fractions were so obtained.

Atomic-weight determinations were made with three of the fractions, the values obtained being 143.2, 185.2, and 225. The latter represented practically pure radium chloride.

Some of the lower fractions of barium chloride that were poor in radium were converted to the bromide and then fractionated. Only one portion, that which should contain the most radium—that is, the last of the four analagous fractions—was saved. The other portions were reconverted to chloride and added to the main chloride crystallization system. In all, 3 grams of pure dry radium chloride and 0.236 gram of radium bromide were obtained from 10,000 kilograms of original residue.

Radcliffe’s Method

Although Haitinger and Ulrich’s method is suitable for the treatment of pitchblende, which contains much more uranium oxide than does the average grade of carnotite, as shipped, the method would not be applicable to the latter ore. The same statement applies to the method that has been used by Radcliffe, as the ore that his method is designed to treat is widely different from the carnotite in Colorado and Utah. The actual mineral carrying the uranium that is found in Olary, South Australia, is carnotite, but it is associated with considerable quantities of ilmenite and rare-earth minerals that are not found in American carnotite.

The dry ore is crushed to pass a 20-mesh sieve, and is then concentrated magnetically; the concentrates amount to about 30 per cent of the original ore, and, as they are insoluble in acids, a fusion process is necessary to effect the initial decomposition. In the fusion process the concentrates are mixed with three times their weight of salt cake and fused in a reverberatory furnace. The fused product is crushed to pass an 8-mesh sieve and is put into wooden vats fitted with agitators. Cold water is fed continuously into the vats at the bottom, and an overflow is provided near the top. By suitable adjustments it is possible to separate out on the bottoms of the vats a considerable amount of comparatively coarse material that is almost free from radium and uranium. The turbid liquid overflowing carries in suspension the radium, lead, and barium as sulphates, together with a considerable amount of finely divided silica. In solution are found the uranium, the rare earths, and part of the iron and “acid earths” contained in the ore. The coarse residues are removed daily, rewashed, and discarded.

The overflow from the dissolving vats is pumped to large lead-lined tanks and allowed to stand all night. The “slimes” settle completely in 12 hours and the clear liquid is drawn off daily and treated for the recovery of uranium. The slimes, which constitute when dry approximately 10 per cent of the weight of the concentrates, are collected weekly and treated for the recovery of radium.

The clear solution, containing the uranium and much of the iron and other bases, together with a large amount of sodium salts, is fed into a series of vats containing a measured excess of a mixture of carbonate and bicarbonate of soda. This is heated and agitated by means of steam jets; the iron and most of the other bases are precipitated, and the uranium, together with some of the rare earths, goes into solution. The iron is filtered off and the uranium solution is made barely acid with sulphuric acid and is heated, the carbon dioxide being expelled by a current of air. The uranium is then precipitated by the addition of ammonia. The ammonium uranate thus formed is thickened in conical settling tanks and in a hydroextractor. The pulp is dried and dehydrated in large muffles. The dried product is broken up and washed repeatedly with hot water to remove sodium salts, the final product containing about 75 per cent U3O8.

To recover the radium from the insoluble residue or slimes from the settling tank, the residue is mixed with half its dry weight of strong sulphuric acid and the liquid is allowed to stand for several days.

The residue is then washed by decantation and by the use of a vacuum filter. This treatment reduces the bulk considerably, dissolving out large amounts of iron salts and “acid earths.” The washed slime, in quantities of about 200 kilograms, is then boiled in large steel boilers with an excess of a 20 per cent solution of sodium carbonate, the solution being replaced once during the boiling. This treatment dissolves a large amount of silica and converts much of the lead, radium, and barium sulphates to carbonates. The slime is then washed until the wash water gives no reaction for sulphates, and is then fed into a warm dilute solution of hydrochloric acid, agitated for a couple of hours, and allowed to settle ail night. The clear solution is siphoned off, and lead, barium, and radium precipitated as sulphates. After having been washed once by decantation, the slime is again treated as above described, two treatments being sufficient to extract most of the radium.

The crude sulphate obtained in this manner is fused with carbonate of soda in large graphite pots and the product digested with hot water. The insoluble residue, after most of the metallic lead has been removed, is thoroughly washed and heated with hydrochloric acid, the solution is evaporated to dryness to dehydrate the silica, and the residue is moistened with acid and digested with hot water, the silica being filtered off.

Fusion with Sodium Carbonate

Another method of radium extraction, that has been used by at least one company, is that of fusing the ore or carnotite concentrates with sodium carbonate; about three times as much carbonate as ore being used. The mixture is strongly heated in a reverberatory furnace lined with magnesite brick, and the fused mass is run directly into vats, in which it is leached. In this manner the silica is converted into sodium silicate and goes into solution together with the uranium and vanadium. The iron, calcium, barium, radium, etc., remain as the insoluble residue, which is washed in filter presses. This material is then treated with dilute sulphate-free hydrochloric acid, which dissolves the carbonates, and the radium and barium are precipitated by the addition of the requisite amount of sulphuric acid or sodium sulphate. The whole is allowed to settle in settling tanks, and the clear liquid drawn off, the barium-radium sulphates, mixed with a considerable amount of silica and other impurities, being left as a sludge at the bottom of the tank. This is taken off without previous filtration and dried, forming a crude radium-barium sulphate, which is then refined by a special process involving fractionation from neutral solution.

Processes Description

As regards these different processes of radium extraction, it can be plainly seen that each has some disadvantages. Any process that involves the use of sulphuric acid as a leaching agent at once puts the radium in an insoluble form. Although, a concentration is usually obtained which may run as high as 10 to 1, or even higher, the advantage of such a concentration is more than overcome by the disadvantage of having to treat the radium as an insoluble rather than a soluble product. The sliming method gives a concentrate consisting largely of calcium sulphate mixed with fine silica, the whole carrying a certain proportion of barium and radium sulphates. Owing to the presence of the silica, none of the short methods for recovering the radium from the insoluble sulphate can be used, and it is necessary to boil the material with sodium carbonate and then leach with chemically pure hydrochloric acid.

The labor required is considerable and makes the cost of refining high. The fact that this method of concentration could be used at the mines was used to justify the extra expense, but as it is cheaper to transport ore than sulphuric acid, the argument does not hold. In addition, the uranium and vanadium concentrate is in an undesirable form which must be retreated before final use. Any process that converts the radium in carnotite into an acid insoluble product, when it is already in the ore in a more easily treated condition, is open to serious criticism.

Leaching with Hydrochloric Acid

Leaching with hydrochloric acid will prove successful with some carnotite ore, provided the acid is pracitically free from sulphuric acid, or contains less than 0.05 per cent of the latter acid, and provided the ore itself is exceedingly low in gypsum and other sulphates. In other words on a selected ore a hydrochloric-acid leach will prove reasonably satisfactory, and an 80 or even a 90 per cent extraction can be obtained under favorable conditions. The preferable concentration is an acid containing about 20 per cent hydrogen chloride, and a weight equal to that of the ore should be used. The filtration must take place while the acid is hot, and, therefore, must be rapid. The residue should be washed with more dilute acid and then with distilled water. Although a satisfactory extraction is frequently obtained, on the other hand the extraction from many ores may be as low as 50 or even 40 per cent, and the method can not, therefore, be considered satisfactory for universal use with carnotite ores.

Fusion with Sodium Carbonate & Acid Leaching

The treatment with sodium carbonate, followed by an acid leach, does not have the same objections. Indeed, this method can be used efficiently for the extraction of radium, the objections being more of a mechanical than of a chemical nature. When the ore is boiled with sodium carbonate, about 80 per cent of the uranium and possibly 60 per cent of the vanadium goes into solution. A previous roast or the use of oxidizing materials during the leaching is advantageous, if not necessary. The sodium carbonate solution can be separated either by the use of a filter press or by settling and decantation, as the liquor does not readily filter by gravity through ordinary filtering media. The ore thus treated must be thoroughly washed with distilled water, in order to remove as much as possible of the sodium silicate and sodium sulphate formed in the reaction. After this washing it is necessary to use chemically pure hydrochloric acid, or at least an acid absolutely free from sulphates, for the presence of small amounts of iron or other impurities is not injurious.

It is almost impossible to filter the acid solution after treatment with hydrochloric acid, because small quantities of sodium silicate remain with the ore, and the liberated silicic acid clogs any filtering medium that may be used. Experiments by the National Radium Institute showed that great difficulty would be experienced in filtering at this stage either by gravity, by suction, or by pressure. Again, settling and decantation might prove successful, but would involve the use of a large bulk of dilute acid. During the time required for settling, the radium would tend to precipitate, especially in the presence of silica. The process also involves the handling of the ore twice, but the main difficulties are in connection with filtration or the separation of the liquors from the residues.

After the acid solution has been removed, the radium can be re-precipitated as radium barium sulphate by adding to the acid solution the required amount of barium chloride and either sulphuric acid or sodium sulphate. The radium barium sulphate can then be obtained by settling and decanting the clear liquid from the precipitate, and finally removing the precipitate through an earthenware filter.

Whether the patent covering this process contains any new discovery might be questioned. Moreover, the same results can be obtained by using sodium bicarbonate instead of sodium carbonate.

Fusion with Sodium Carbonate

The fusion methods used for pitchblende and the Austrian ores do not apply to American carnotite, although they may have some advantages for the particular ore they are designed to treat. The fusion of carnotite with sodium carbonate has some advantages and some serious disadvantages. The main advantage is that it is adapted to carnotite in any form whether it be ore of 20 or 30 mesh or concentrates of more than 200 mesh. In addition the presence of sulphates in the ore is not deleterious, as they are removed at the same time as the sodium silicate. The disadvantages, however, are numerous. In the first place, as a large bulk of sodium carbonate has to be added to the ore considerable material has to be handled. After the fusion 3.5 tons of fused material are handled for 1 ton of ore or concentrates. The uranium and vanadium is in the filtrate from the leached material, and the presence of so much sodium carbonate and sodium silicate makes the cost of the recovery of the uranium and vanadium excessive.

The uranium can be recovered in one or two ways—either by adding sodium hydroxide directly to the filtrate, or by making the filtrate slightly acid with sulphuric acid and then adding sodium hydroxide to the hot liquid. In the first case the amount of sodium hydroxide required to precipitate the uranium in the presence of so much sodium carbonate is large and usually involves the precipitation also of a considerable amount of silica. The result can be more easily accomplished by first making the solution acid, but this involves again the use of a large amount of sulphuric acid. In addition the crude barium sulphates obtained are mixed with a considerable amount of silica, which makes them difficult to treat by ordinary methods.

It is doubtful whether a total recovery of radium of more than 70 per cent has been obtained by this process. This low recovery and the necessarily high costs more than counterbalances the advantages that the method may have.

Possible New Methods of Radium Extraction

In the search for a better method than those outlined the main object to be borne in mind is a high extraction and recovery of the radium. The importance of a high recovery of uranium and vanadium has been, to some degree, unduly emphasized; the main object is to obtain the radium.

Advantages of Using Nitric Acid

For extracting the radium the use of nitric acid possesses many advantages, especially if the initial cost of the nitric acid can be reduced through the recovery of sodium nitrate as a by-product. The cost of the treatment would be very largely increased if 6 to 7 cents per pound of 100 per cent nitric acid had to be paid instead of the lower cost of hydrochloric acid. On the other hand, if, in connection with the nitric acid method, sodium nitrate can be recovered with small loss and nitric acid be once more made from the sodium nitrate a cycle would be obtained that would reduce the cost of the nitric acid to a figure as low as, if not lower than, that for the hydrochloric acid.

In addition, the solvent action of nitric acid on radium sulphate is much greater than that of hydrochloric acid. Radium belongs to the calcium, strontium, and barium group, and, as analysts know, in order to precipitate barium sulphate completely and efficiently free nitric acid must be removed. Consequently, nitric acid has a much greater solvent action on radium even though the latter may be associated in the ore with reasonable quantities of sulphates.

This greater solvent action of nitric acid was the general basis for the method devised by the Bureau of Mines and used in the plant of the National Radium Institute. It has recently been recognized by Plum, who suggests that possibly the best method for treating carnotite is to boil the ore first with sodium carbonate, leach the washed residue with hydrochloric acid, and follow this leaching with nitric acid in order to dissolve the 10 per cent of radium that he was unable to extract with hydrochloric acid. Plum had in mind the extraction of both radium and other radioactive constituents, such as polonium and actinium, but he plainly indicates that nitric acid can dissolve out of carnotite radium that can not be dissolved by hydrochloric acid, even after the ore has received a preliminary leaching with sodium carbonate. This finding was also established in the preliminary cooperative work carried on two years ago which led to the adoption of the nitric acid method.

When the use of nitric acid at the start will give just as high an extraction as the combined use of sodium carbonate, hydrochloric acid, and nitric acid, there is no need to use the three steps unless it is desired to recover the other radioactive constituents in the ore.

By the method he outlined, Plum was able to recover 89.9 per cent of the radium in the ore on a laboratory scale, using 1 kilogram of material; this extraction has been exceeded a number of times on carload lots in the plant of the National Radium Institute by the use of nitric acid alone. As a scientific accomplishment the recovery of the polonium, actinium, and ionium would be strongly advisable, but as these now have little or no commercial value they can not be taken into consideration in a commercial process.

In the nitric-acid process the radium is at once recovered as a high-grade radium barium sulphate, practically free from silica, and easily treated by improved methods. The process is adapted to recovering either the radium by itself, or the radium, uranium, and vanadium, as the radium is obtained first, and from that point all the other products may be discarded without further treatment, if so desired.

The different steps are each completed in one day, the equipment is not expensive, the extraction and recovery are high, and it is believed that the costs are lower than those with any other process for treating carnotite.

From the data already presented it can be readily seen that each particular type of radium-bearing ore has to receive a more or less different treatment, depending upon the other constituents of the ore. This process has been tried only with American carnotite, and may not be adapted to other radium-bearing ores, as its efficiency in connection with other ores has not yet been fully determined. The process is, however, applicable to the treatment of carnotite ores obtained in Colorado and Utah.